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By Donald C. Craighead

"Overland conveyor systems to transport bulk materials from mines to processing plants or load-out stations have to be seriously considered as the viable alternative to truck and train haulage systems. With the ever-increasing price of diesel fuel, conveyors with their electrical drives generally prove to be the least costly when operating and maintenance expenses are analyzed for the life of the installation.IntroductionOverland conveyor systems are now considered as one of the viable alternatives which should be closely investigated when a mining company and its engineers initiate studies on methods of transporting raw materials from a mine to a processing plant or rail load-out station.This paper will generalize the overland conveyor and compare it with alternative handling systems. Each installation must be carefully studied to determine the best route, type of foundations river crossings; amount of excavation work for cut and fill, maintenance roads, etc. No two locations are the same nor are the principles of operation of an overland conveyor the same as those of a normal plant or mine belt conveyor. The proper time and effort must be put into the study of the system to evaluate the costs for comparison with other systems.The length of the overland conveyor depends on the layout of the mine and plant and the topography of the terrain between the terminals. Individual lengths may be up to 20 km, and a series of conveyors may extend 100 km or more.Advantages of Overland ConveyorsThe main advantage of overland conveyors is the low operating and maintenance costs which will be realized over its years of operation. Although truck haulage probably will permit lower capital investment during the initial phases of the mine development when the annual capacities have not developed to the design level, their operating and maintenance costs will be considerably higher. When evaluating the total costs per ton of material to be handled during the expected life of the operation and the projected increases for diesel fuel versus electricity, overland conveyor systems will generally prove to be economically more viable."

Jan 1, 1982

By John Drew Ridge

For the early conservationists before the first decade of the twentieth century, conservation meant largely the planned preservation of water, forest, soil, and wildlife resources-renewable natural resources-through efficient use. Water was primarily thought of as a resource that must be kept available for humans to enjoy in a recreational sense and as a habitat for wild animals, birds, and fish. Little thought was given to conserving water for actual human or industrial needs. If industry entered the picture it was as an actual or potential polluter of water, which was the basic necessity for drinking and bathing and for outdoor recreation. Even the forests were only held worthy of conservation largely because they were areas in which wildlife could, or should, abound and in which human beings might find relief from the cares and problems of everyday life. Finally, the soil was to be preserved in its natural fertility through proper fertilization, cultivation, and crop rotation; in this the conservation program did enter the realm of the practical. For through soil conservation, the world was to be able to raise increasingly large crops of constantly improved organic foods which the steadily growing population of the earth was to need to survive and improve. Relatively little time, however, was spent on considering what profit the farmers of the world were to make from the scientific farming on which they were to embark. In short, the appeal of the early conservationist program was much the same as that of the entire era of Theodore Roosevelt of which it was a part; it was to the heart rather than to the mind; it was to emotion rather than to reason. Because of the way in which the problem was approached, conservation took on the character of a crusade rather than that of a practical plan for the betterment of man through economically sound methods of improving his environment and the uses he makes of it. Conservation had, then, all the advantages of a stirring appeal to the emotions, but it did not have the lasting results that would have been achieved by a program with a sound economic basis.

Jan 1, 1959

By J. M. Callow

THe results obtained by pneumatic flotation throughout the country on all classes of ore, and the tonnage now being treated by this particular method, speak for themselves. Its advantages over the so-called agitation process are now established by facts and data from practicc. The comparative efficiency of a rapidly revolving paddle, as a means for aerating an ore pulp, and of a cyeloidal or turbine blower for the same purpose, would seem to be so obvious as to leave no room for discussion. Practitioners in this art are to be congratulated on the acquisition of the valuable literary contributions that have recently been made on the theory of flotation by Messrs. Ralston, Van Arsdale, Taggart and Beach, and others, and the very valuable and able articles by Messrs. Laist and Wiggin on "Flotation Concentration at Anaconda, Mont.," by Dr. Rudolf Gahl on the "History of Flotation at Inspiration," and by David Cole on "The Advent of Flotation in the Morenci District." It is not within the province of this article to attempt an answer to the various issues raised in regard to the theory, or to defend the one advanced by me a year ago. But since discussion usually benefits the profession, a reply will be made to some of the statements made by the last-named writers. CoMPABATIVE SENSITIVENESS AND ATTENDANCE Concerning the article, "Flotation Concentration at Anaconda, Mont.," by Laist and Wiggin, a little explanation is in order. It should be stated that, with the exception of about 24 days at the end of the testing period, the Callow experimental plant at Anaconda was not run under my direction. This perhaps accounts for some of the concIusions drawn on page 499 of the descriptive paper. The conclusion that the use of acid seems to be of considerable advantage, calls for some comment. While the tests conducted by the Anaconda Copper Co. do show acid to be of considerable benefit, this is a consequence of using acid sludge as a flotation agent. My belief has always been that the Anaconda

Jan 1, 1917

By DeVerle P. Harris

Risk for most mineral investments is greater than that of average business ventures, because the physical and economic characteristics of the mineral deposit are never known with certainty. Inasmuch as the merits of a potential mineral investment cannot be fully appraised without consideration of the impact to the investor of the possible unfavorable events associated with the adoption of the project, methods of risk analysis are important in project evaluation and investment selection. The conceptual framework for risk analysis and the means by which some of the recommended practices might be implemented are examined. In addition to the normal business risk due to market forces, the physical and economic characteristics of the mineral deposit are never known with certainty, for only partial or indirect information is available, such as data from drill cores or surface and underground workings. Consequently, risk for most mineral investments is greater than that of average business ventures. Inasmuch as the merits of a potential mineral investment cannot be fully appraised without consideration of the impact to the investor of the possible unfavorable events associated with adoption of the project, methods of risk analysis are important in project evaluation and investment selection. This paper examines the conceptual framework for risk analysis and means by which might be implemented some of the recommended procedures. Utility Theory A discourse on risk and its impact upon investment selection quite often leads the reader into the concept of utility theory, for utility theory is a device constructed by economists to measure the satisfaction an individuai receives from a good; in the case of investment analysis, that good is the outcome of an investment. It briefly is reviewed here to describe the reaction of the individual to risk and as a stepping stone to other techniques more suited to risk analysis at the level of the firm. Utility theory proposes that an investor possess a utility map (curve), such as indicated in Fig. 1, which indicates the value in utiles of potential monetary rewards to an individual with some specific endowment of resources. Points on this map are obtained from the following relationship:

Jan 1, 1971

By Thomas M. Price

Sand production for use in hydraulic concrete is probably the most important element affecting the physical characteristics of the resulting concrete in the opinion of most engineers, aggregate producers and contractors of today, who are engaged in the effort to produce stronger concrete per unit of cement used. Their opinion is based on both laboratory and field results. In the past, proper cement production has been the object of most of the study given this subject and while it is still studied by the extensive and well equipped laboratories that are a part of every cement mill, as well as by all governmental laboratories, at the present time the object of most of these studies has been in the field of low heat and quick setting qualities rather than the effect on the strength of the concrete. It is recognized that concrete must have two important qualities; namely, workability with the least water possible, and uniformity. These qualities have been accepted as axiomatic and possible to obtain if the foreman in charge knew his business. Now, however, with concrete produced by plants that mechanically weigh and chart all the ingredients, including water, of a batch of concrete, the foreman cannot cut and fit his aggregates, cement and water to fit his own desires and the secret of the important qualities of workability and uniformity are found to lie largely in the sand, hence the attention given its production at the present time. With this condition in mind the engineers of the United States Bureau of Reclamation and the Six Companies Incorporated in designing the plant for producing the aggregates for the Boulder Canyon Project included every possible means for the production of a sand of a constant and acceptable fineness modulus as well as a constant moisture content, and for attaining these results with as few detrimental ingredients, like silt and vegetable matter, as possible. This aggregate plant (Figs. 1 and 2) is in Hemenway Wash, 7 miles north of Boulder City. It is 4 miles upstream from the low-level mixing plant. At this screening plant 8,234,000 tons of raw aggregate are being

Jan 1, 1934

By George Rice

THE subject of dust explosions in coal mines first appears in the Transactions of this Institute following the first great mine disaster that happened in bituminous mines of the United States. This was the Pocahontas explosion, which occurred Mar. 13, 1884, causing the loss of 112 men. The operating company wisely asked the technical assistance of the Institute, requesting that a committee be appointed to investigate and report. This committee consisted of Messrs. Bramwell, Buck, and Williams, who, after careful study, reached the conclusion that the explosion was due "either to dust alone or to dust quickened by an admixture of fire-clamp too slight for detection by ordinary means." This is perhaps the first time in this country that coal dust was considered to be the chief agency in a mine explosion.. The report is given in the Transactions, vol. xiii, p. 237 (1885). It is particularly worthy of note that the committee called attention to "the necessity of full-sized tests and systematic experiments. " In the same volume of the Transactions, there is a most interesting paper reviewing the literature relating to coal dust, by E. S. Hutchinson, entitled Notes on Coal-Dust in Colliery Explosions. Mr. Hutchinson gives a historical account of observations on the part played by coal dust in mine explosions, beginning with Messrs. Lyell and Faraday's report on the Haswell colliery explosion in 1844, in which it is said "much coal gas had been made from this dust in the very air itself in the mine by the flame of the firedamp, which raised and swept it along, and much of the carbon of the dust remained unburned only from want of air." The subject lay dormant until 1875 when William Galloway began some coal-dust experiments in a small gallery at a mine in South Wales of which he was manager. At that time he considered the presence of firedamp necessary for continued propagation of a dust explosion; but in 1884, the year of the Pocahontas disaster, he stated that no earlier author-than himself had credited coal dust with being a principal factor in mine explosions, relegating firedamp to a secondary place.

Jan 10, 1914

By A. I. Krynitzky, Henry S. Rawdon

The most commonly used material for terminals in spark plugs is commercial nickel wire, because of its relatively high temperature of melting, excellent heat conductivity, and slow rate at which the metal is oxidized, even upon continued heating at high temperatures. The grade commonly used for this purpose averages 97 per cent. nickel, the remainder being manganese, cobalt, iron, copper, with traces of other impurities always found in commercial nickel. A peculiar and interesting type of deterioration that occurs in these nickel terminals during the service life of the spark plug was recently brought to the attention of the Bureau of Standards. Commercial spark plugs vary greatly in their size and shape and the terminals differ considerably as to their relative size, shape, number, and arrangement with respect to one another. Although this defect was studied in detail in only one form of spark plug, the deterioration is a characteristic of the material, i.e., of nickel wire, rather than of the particular type of spark plug in which it was noted and studied. The mechanical features of the plugs, however, have a considerable bearing on the time required for the deterioration to become serious. The results of this study of the deterioration of nickel terminals in service should be of value to the makers of all types of spark plugs in which the terminals are made of this metal. CharacteR of Deterioration of WIres During Service Macroscopic Appearance and Properties of Changed Wire.—The nickel terminals in which the deterioration was studied were taken from spark plugs having a central terminal with the ground or side terminals attached firmly at both ends to the shell of the plug. In some of the plugs, two side terminals were used, one on each side of the central one. Although both terminals were found to have deteriorated to some extent, the attack of the central one was quite negligible compared to that of the side ones. These latter wires had developed, in service, transverse cracks that, in many cases, were as sharp and definite as a knife-cut. After a separation occurred, the breach widened by loss of material from the

Jan 1, 1921

By Charles Fulton

EXPERIMENTAL work on the process was begun on a laboratory scale at Cleveland, Ohio, in 1914, and transferred to East St. Louis, Ill. in 1916, where a commercial sized furnace was in technical operation until January, 1918. The essential steps of the process are as follows: Oxidized zinc ore or roasted zinc concentrate is mixed with crushed coke and coal-tar pitch and formed into briquets 9.25 in. (23.5 cm.) in diameter and 21 in. (53 cm.) long in a manner similar to that used in the manufacture of graphite or carbon electrodes, except that much less care and time are required. The composition varies with the nature of the ore; in one case, the composition was 100 parts ore, 70 parts coke, and 18 to 20 parts pitch. These briquets maintain their form and volume during and after the distillation of the zinc. This object is gained by using coke as the matrix and coal-tar pitch as the binder; this pitch becomes coke on heating and unites the ore particle and the original coke particle into a continuous mass. The briquet is an electrical conductor but only to such, a degree that it can be used as a resistor. The size of the ore, as in the present practice of zinc metallurgy, ranges from fine material, such as flotation concentrate, to coarse table and fine jig concentrate. The process is not restricted to any particular kind of ore, but is applicable to pure high-grade ores, ores high in iron, and complex zinc lead ores, since the residue from the distillation is held immovably in place within the briquet. In the case of complex zinc-lead ores, the distillation is conducted with a carefully regulated temperature so that a high percentage of lead is retained in the distilled briquet, which may then be smelted for its lead content. In present zinc smelting practice, the reduction fuel is from 40 to 50 per cent. of the weight of the ore. In the briquet, it is from 60 to 85 per cent. of the ore; but unless the ore is exceptionally high in residue, the distilled briquet

Jan 9, 1919

By S. D. Strauss

The marketing of nonferrous metals and of the ores and concentrates from which these metals are recovered is a fascinating trade, international in character, sensitive to every change in the economic tide, and complex in its organization. Each individual metal has its own marketing mechanism; no two are exactly alike. The differences sometimes are a matter of tradition rather than of necessity, but usually they have developed because of the very nature of the problems of producing or consuming the individual metal. This description, because of limitations of space, can do no more than indicate the major aspects of the trade and touch lightly on the marketing structure of each metal. The dictionary states that a market is a "meeting of people for selling and buying." In the case of nonferrous metals this meeting may occur on a half-dozen different occasions between the time the ore has been first mined and the time the refined metal is available for use by industry; or it may not occur at all until after the metal itself has been rolled, drawn, stamped, cast, or forged into a useful semifabricated shape. To illustrate: A Chilean miner may sell his small ore production to a custom mill, which produces concentrates that are sold to an ore dealer, who finances the shipments of the concentrates to a United States custom smelter, which buys these' concentrates for conversion into blister, which is subsequently refined into electrolytic copper and sold. Conversely, much of the production of aluminum is in the hands of integrated concerns that mine bauxite, convert bauxite to alumina, smelt the alumina to produce aluminum, and then fabricate the aluminum into semifinished or even finished products before title to the material changes hands for the first time. Some nonferrous ores are not converted into metal at all but are sold by the miners or dealers directly b the ultimate consumers. A good example would be chrome ores used by the chemical industry. And of course a substantial part of the commerce in nonferrous metals is in scrap, which is the reconversion into useful form of metals that have long since been mined and smelted. Thus, one-third of all the lead used in the United

Jan 1, 1959

By Dr. O’Neil Thomas J., Donald W. Gentry

There are mines that make us happy, There are mines that make us blue, There are mines that steal away the tear-drops As the sunbeams steal away the dew. There are mines that have lost the ore chute faulted, When the ore's forever lost to view, But the mines that fill my heart with sunshine, Are the mines that I sold to you. -Anonymous INTRODUCTION The preceding chapters have discussed, in various degrees of detail, the primary considerations necessary to perform a cash flow analysis of a mining-related investment opportunity. Although the previous discussions regarding the manipulations, calculations, and interrelationships of the various components of the pro forma income statement were illustrated in the text, many individuals are un- familiar with the format and appearance of a final spread sheet for the evaluation of a mining property. As a result, it seems appropriate to conclude this book with an abbreviated case study of an actual mine property (currently in production) showing the cash flow analysis and some back-up data. Obviously all of the supporting data and calculations for the case study presented cannot be provided; however, reference is made to the major assumptions utilized and to the appropriate procedures employed in the calculations, as described elsewhere in the text. The problem associated with using any case study for illustrative purposes is that one tends to be overwhelmed by detail. For instance, specific costs, prices, expenditure timings, and other variable values may be challenged because they either do not conform to one's expectations or are not consistent with specific operating experiences. When this occurs the instructional value of the problem becomes obscured. The purpose of most case studies, and particularly the one presented in this chapter, is to expose the student to the methodology, procedure, and format associated with performing a cash flow analysis of a mining property. It is not to laboriously wade through the specific details associated with operating and capital cost development, mine plans, or equipment schedules. It should be noted, how- ever, that the values represented in this case study result from specific mine plans, operating procedures, equipment spreads, manpower requirements, and other en-

Jan 1, 1984

By A. M. Croll

The efforts to mechanize narrow reef tabular orebodies unique to this country, are not new. They began in the early 1900s and have over the last century seen varied rates of progress and success. Efforts to mechanize with the use of scraper winches can be regarded as a success. The scraper winch replaced the hand lashing of ore and development waste into hoppers. The initiatives by Anglo Gold and the then JCI to implement trackless mining equipment into these environments had limited success but in general were looked upon as a failure?at least this is the perception. The Chamber of Mines Research Organization (COMRO) in conjunction with various mining houses tried to make a success of the impact ripper. Diamond wire cutting of the stope face is another, which has been tried and tested. The use of high-pressure water to cut rock has also been trialed. These attempts, amongst others, are just a few of the efforts that have not been successful in the quest to make life at the face safer, more productive and easier for our employees. Against the backdrop of an emerging South Africa, the associated high expectations of historically disadvantaged people within the country, a strong rand, demands for better safety, lower cost of production and more efficient operations, mechanization has the chance to make a difference to Anglo Platinum?s business. Mechanization has certainly changed the way other commodities such as coal, base metals and diamonds are won from the earth. Mining?s image is still one of people crawling around in stopes?a dirty, ?low-tech? business. Anglo Platinum?s latest journey to mechanize its underground operations began in earnest in1998 and has enjoyed some successes and failures. The opportunity to change the way mining is done within the company is there. Technologies are evolving, and the implementation of mechanization can certainly assist with the achievement of the goals referred to above. When mechanization commenced, targets were set for the company to achieve but these have been revised in the light of recent events relating to the strengthening rand and a slow-down in expansion plans. However, the intent to mechanize appropriately in Anglo Platinum still exists.

Jan 1, 2004

By David Miller

"The biggest concern about using soil mixing is the inability to observe physically the final product, which has led to a lack of confidence about what is in the ground. After all, you can’t see it and you can’t taste it and you can’t smell it, so are you sure it’s soil mixing? In addition, there are variations in techniques, tools and monitoring that have led to additional skepticism and the desire to “elevate” the product. This skepticism is why the industry is so scrutinized even when there are other technologies being used that do not experience nearly the same level of scrutiny. With other techniques, even with a slurry wall, you can see the backfill before it goes into the ground, and drilled piers have careful evaluation of the concrete mix even before placement into and verification of the open hole. Essentially, you get control over the final material. There is less control in soil mixing, and it’s very dependent on in-situ conditions. These factors make it difficult to rely solely upon installation parameters, quality control (QC) and verification tests. At some point, you may need to apply some engineering judgment.Start at the BeginningThe availability of sufficient and accurate geotechnical data (soil type, heterogeneity, shear strength parameters, variability, etc.) is important in establishing the material that will be drilled. This data will also help in determining the proper mix design to achieve the desired soil cement properties. Many factors must be considered when designing the mix and the mixing program, including any debris in the ground, different soil properties and varying moisture contents.Each of the various factors affect the final properties of the soil cement, so it really takes careful planning and evaluation of the mix design through iterations by looking at the soil types and adjusting the mixes. Clay is going to behave very differently than sand in both mixing efficiency and final properties. That is, the presence of clays results in more difficulty in achieving higher strengths and to physically mix the soil and the binder, whereas the presence of sands results in more difficulty controlling the final permeability."

Jan 1, 2017

By Francis Lucas

A microscope objective of given numerical aperture, whew used with light of given wave length, has some fixed limit of resolution. This may be expressed as potential resolving ability-the ability to resolve is inherent in the lens but whether this limit is achieved in practice is quite another matter. Nearly 50 years ago Abbe evolved theoretical considerations which indicated that the ability of a lens to resolve detail is directly proportional to twice the numerical aperture of the objective and inversely proportional to the wave length of the light used. It is true, unquestionably, that as the numerical aperture of the objective is increased the potential resolving ability of the lens is increased and likewise whew the wave length of the light used is decreased, the degree of resolution is increased. Whether Abbe's formula expresses the exact relationship is not a matter of great concern. It suffices to know that to improve resolution, other things being equal, two avenues of approach are available. Either the numerical aperture of the objective may be increased or the wave length of the light decreased. With the apochromatic system it is possible to employ successfully wave lengths from 400 to 500, µµ and still secure reasonably short exposures. The present limit for numerical aperture is 1.40 N. A. although objectives of 1.65 N. A. are wow an experimental possibility for metallurgical work. Thus certain well defined limits are set for those who work within the range of the visible spectrum. In order to take advantage of the increase in resolving power which accompanies a decrease in the wave length of light employed, Dr. A. Köhler and Dr. M. Vow Rohr of the Zeiss Scientific Staff developed the "monochromats." These are quartz objectives corrected for the single wave length 275 µµ which lies in the ultra-violet region of the spectrum. They also developed a series of quartz oculars and a complete photo-micrographic equipment with which to use the quartz optics. Quartz must be used for the optical parts because glass absorbs ultra-violet light almost completely. Thus there was made available about 20 years

Jan 6, 1926

R. Schuhmann, Jr. (Purdue University)— Fulton and Chipman's results on rate of silica reduction from slags are analogous in many was to the results of Parlee, Seagle, and Schuhmann10 on rate of alumina reduction from alumina crucibles. Both investigations have given comparably low rates of reduction and slow approaches to equilibrium. Accordingly, we may hypothesize that the rate-determining step is the same in both kinds of experiments; that is, oxygen diffusion across the stagnant boundary layer on the liquid-metal side of the interface between the liquid metal and the oxide phase (slag or solid oxide). I suggest that silica reduction involves the following consecutive steps: I) At the slag-metal interface: SiO2(slag) Si+ 20 II) Transport of oxygen from slag-metal to gas-metal interface: a) diffusion across liquid-metal boundary layer at slag-metal interface. b) convection within the body of liquid metal. c) diffusion across boundary layer at metal-gas interface. 111) At the metal-gas interface: C +O- CO (gas) Iv) At the graphite-metal interface: C (graphite) -C At steelmaking temperatures it is reasonable to assume that equilibrium is attained in all three chemical reactions (I, 111, and IV) right at the respective interfaces. Convection within the stirred liquid metal (step IIb) is also rapid. Transport of Si and C should be orders of magnitude easier than transport of 0, because of the relatively high concentrations of Si and C. Accordingly, we might expect the overall reaction rate to be determined by boundary-layer diffusion of oxygen, either IIa or IIc. Fulton and Chipman's demonstration that bubbling CO through the system had no major effect on reaction rate indicates that IIc is not the slowest step. Therefore, it becomes logical to estimate the maximum rate for step IIa and to compare this theoretical estimate with Fulton and Chipman's experimental data. If oxygen diffusion across the liquid metal boundary layer at the slag metal interface (step IIa) is rate-determining, In this equation, dn sio, /dt is the rate of silica reduction in moles per sec,A is the area of slag-metal interface in sq cm, Do is the diffusivity of oxygen in sq cm per sec, 6, is the boundary layer thickness in cm, c,* is the oxygen concentration right at the slag-metal interface in moles per cubic cm, and co is the oxygen concentration in the body of the liquid metal, also in moles per cubic cm. Equilibrium data" on the silicon deoxidation reaction in liquid iron and steel at 1600°C indicate that the oxygen contents of the liquid metal in Fulton and Chipman's experiments at 1600°C probably fell in the range of 0.5 x10-3 x10-3wt pct. That is, the maximum conceivable value of co*-co for the system under consideration was on the order of 10"5 mole oxygen per cubic cm. On the basis of previously published data,1O,11 it is estimated that Do/0 will fall somewhere in the range from 10-3 to 10-1 cm per sec. The surface area A in Fulton and Chipman's experiments was approximately 20 sq cm, and the weight of metal involved was about 500 grams. Combination of all these figures with the above rate equation leads to an estimate that the rate of silica reduction should fall within the range from 0.002 to 0.2 wt pct Si per hr. This estimate is consistent with the experimental data. For example, Fulton and Chipman's Fig. 2 shows a change of about 0.3 pct Si in 10 hr, corresponding to an average rate of 0.03 pct per hr. According to the proposed hypothesis, increasing the temperature will increase the reaction rate ill two ways: 1) by increasing oxygen diffusivity and 2) by increasing the oxygen concentration (oxygen solubility) in the liquid metal. The combination of these two effects accounts for the high value of the observed activation energy. Summarizing, I propose that the rate of silica reduction, like that of the carbon-oxygen reaction, is diffusion controlled and that low oxygen concentration in the liquid metal is the major factor accounting for the very low observed rates of silica reduction. John Chipman (author's reply)—The authors thank Professor Schuhmann for his interesting contribution. His proposed explanation may well prove to be the correct one. There is clearly a need for much further experimental work on this problem, and further research is in progress.

Jan 1, 1961

CHAIRMAN LITTLEJOHN: Some of the members will want to question Mr. Farnham on some of the things he has brought tip, so we will throw the meeting open for discussion. I am quite sure Mr. Farnham will be willing and glad to answer any questions that may be put to him. MR. HOLMAN: I would like to ask Mr. Farnham a question in regard to the Duckbill: How wide a face will it operate without moving the bill? MR. FARNHAM: The Duckbill is a device that was developed in the Union Pacific Mine at Rock Springs, within the last few months, and it is shaped something like, a dustpan, with discharge throat in the rear. They made one seven feet wide and found it was too wide, and the one they are using now is between three and four feet in width. The front section of the conveyor, which attaches to the duckbill, meshes into it to form a sliding section, and the front end of the Duckbill is pushed by the upper segments. I saw one used in a room twenty-four feet wide, and they drew it back after they cleaned out one room and pushed it forward again. MR. HOLMAN: You have no arrangement other than ratchet work? MR. FARNHAM: The ratchet is for forward working and back; the sections are made 12 or 13 feet long, and the stroke of the conveyor automatically works the ratchet; they can ratchet it back; it is a hand device. MR. HOLMAN: They just swing the end of the conveyor across the face to pick up the loose coal? MR. FARNHAM: Yes; they are also planning to use it endwise along the panel face, that and the room work both are quite new in their application, and I do not think the Union Pacific is giving out any results as yet, because the system has not been in use long. MR. HOLMAN: We are planning on putting one in very shortly. MR. HALL: I understand they can swing it over 30 degrees one way or the other, and it goes the same way easily because of the continual jar of the back and forward motion on a 60-degree swing. SECRETARY SHUBART: Take on a comparatively long conveyor, you can get quite a sweep, a large part of the circle. The Duckbill could command quite a large range. For a 12-foot entry, take a four-foot wide advance through the middle of the shot coal, with a 12-foot advance, ratchet back, swing the Duckbill to the rib and repeat; but I do not think they could get as much as 60 degree. on the end of the Duckbill without curving further back. The conveyor is light and not normally set on any permanent foundation. I saw one of Mr. McCarty's first machines, a rather rudimentary one, clean tip a 12-trot entry, 21 tons of coal, in 16 minutes, three men at the face and one at the loading end. MR. FARNHAM: The average time to clean the entries is about an hour-to clean up good-it wouldn't take but a short time; that is why they get this record of several cuts a shift. CHAIRMAN LITTLEJOHN: Are there any other questions you would like to ask Mr. Farnham? MR. HOLMAN: We were discussing this entry driving in regard to the entry loader; an entry loader in advance work, just what you have to do to place the jackbars with the sheave-blocks? Do you dig along the ribs after the coal is shot or do you have any other scheme that could be worked out to a good advantage? MR. FARNHAM: This record is a report I have just received from our Chicago office, without the full details, which I have written for. I am sorry I cannot give you the details. This was an Indiana or Pennsylvania mine, near Pocahontas, not the one where they had an accident. I assume they set the jacks ahead and swung and dumped the coal partly out with a scoop and partly by hand, but the application for that for entry driving is entirely new. The reports stated that they anticipated using that as a standard method of entry driving in the future. I will be glad to see that you get a copy of it when I get the report completed. MR. MONAY: Mr. Shubart, did you say they loaded 12 tons in 16 minutes? SECRETARY SHUBART: Twenty-one tons. MR. MONAY: How long did it take to change the machinery out of there

Jan 1, 1926

By J. L. Smith, A. L. Albee

This article presents a series of guidelines by which the geologist can evaluate the likelihood of surface faulting and its probable extent at any given site in Southern California and Nevada. The information is intended primarily for geologists concerned with establishing design criteria for proposed nuclear plants. The geologic problems involved in the location of a nuclear power plant are fundamentally no different from those for other types of installations. They fall into four main categories: foundation stability, landslides and slope stability, shaking due to earthquakes, and surface faulting. For problems in the first three categories the foundation engineer, the geologist, and the seismologist can provide criteria for plant location and design, and these problems can generally be economically handled by appropriate design measures for a project of the magnitude of a nuclear plant. However, problems in the last category — surface faulting — are more difficult to handle and require a unique evaluation. Accordingly, this paper will deal primarily with the problem of establishing design criteria for surface faulting, particularly as it affects nuclear facilities. A nuclear reactor is a power source that for greater safety is contained in a heavy, air-tight structure, just as gas, oil, water, and other power sources must be contained. Surface faulting is significant in that it may reduce the integrity of the containment by affecting critical exterior piping or by breaching of the containment. A similar significance exists relative to dams or tanks for the storage of water, gas, or oil, except that in these latter examples the breaching of the container automatically releases the fluids to do their damage. This is not necessarily the case with the rupture of a reactor containment structure because the function of the containment is totally protective, i.e., it is necessary only in the event that radioactive products are released from the reactor, and there are many other safeguards to pre- vent this release even if the containment is ruptured. At the present time, nuclear power plants must be located near large sources of water for cooling the steam generated. The construction of an industrial facility of any kind on the coast line is esthetically distasteful to most people since, unfortunately, there is not enough coast to fulfill all the needs and all the desires of all the people. In most cases where industrial facilities encroach on the lives of citizens, there is no mechanism other than zoning laws by which a person can effectively protest. But in the case of a nuclear facility, the public hearing required by the Atomic Energy Commission provides a forum for dissent, as in recent case histories, and the question of safety provides objectors with a weapon for fighting the construction of the plant. The nature of a public hearing for a nuclear plant is such that the prospective owner and operator must prove that there is no undue hazard, whereas the objectors need only demonstrate that there is a reasonable doubt. It is in this situation that geology becomes the Achilles Heel of nuclear power plant location. For his investigation, the geologist has natural exposures of rock at the ground surface and a limited number of trenches and drill holes to give him a fairly complete picture of the distribution of the various rock types. From the surficial data, he must infer three more dimensions — depth below the surface, past time, and future time. Unlike many problems faced by engineers, the geologist has only this one set of data from which to reach a conclusion — he is unable to reproduce the natural sequence of events in order to obtain another set of data for comparison. Hence, by the very deductive nature of a geologic conclusion, it is difficult to prove a geologic conclusion beyond a reasonable doubt even to other geologists — and perhaps one should say especially to other geologists because the experience and background of a geologist will strongly influence his conclusion, and no two geologists have exactly the same experience and training. The engineer and the public official would like the geologist to conclude that faulting cannot occur at a given site or to assign a numerical value to the probability of its occurrence — but no responsible geologist can do either of these things. Since government officials and others must make decisions that affect the public safety, it would seem that the geologic profession must attempt to establish criteria and

Jan 1, 1968

By W. E. Rudolph, R. E. Baylor

DUE to its location in the Atacama Desert, one of the most barren of the earth's surfaces, Chuquicamata's water supply presents unusual problems. Yearly rain-fall averages less than one tenth of an inch at the plant. However, there are summer showers above 12,000 ft in the Cordillera to the east, the resulting run-off flowing through old river valleys buried beneath more recent volcanic formations, to be impounded within sediment-filled basins. This water emerges at springs where the outlets of these basins are blocked by lava flows, and here are formed the small streams which feed the only important river of the region, the Rio Loa. Chuquicamata's water is obtained from these springs and rivulets. [ ] The map above indicates four pipe lines from which potable and industrial water are supplied. Potable water, amounting to 4500 metric tons per day, is conveyed in the Toconce pipe line from springs 59 miles due east of Chuquicamata. This water is used not only for drinking, but also for boilers and other needs requiring high quality. For industrial water at the oxide plant, there are two 12-in. pipe lines from the Rio San Pedro, carrying a total of 17,000 metric tons per day of slightly brackish water. This water is at present used mainly for leaching and for hygienic purposes. Water Source Found For the present and future needs of the sulphide plant, it was calculated that at least 32,000 metric tons per day of make-up water would be required. For this purpose, a pipe line of 44 miles length was constructed to bring in the entire flow of the Arroyo Salado, one of the eastern tributaries of the Loa. The salt content of this water is so high (over 5000 parts per million of solubles, mostly chlorides) that it is highly detrimental to farming, and the Chilean Government had been studying projects to separate these waters from others of the Loa system in order to improve agricultural conditions in the fertile valley of Calama. So it happened that the Government was willing to award rights to the Arroyo Salado waters under agreement whereby the Mining Company removes waters from the Rio Loa system above Calama for all time. The outlet of these waters, after serving their purpose at the new concentrator and leaving the plant in tailing, is the Salar de Talabre, an old salt lake which presents fully ten square miles of surface to serve as an evaporating pan, the outlets having now been blocked by dams. Here the dry climate of Chuquicamata is a favourable factor, evaporation averaging slightly above 1/4 in. per day. The Toconce and San Pedro pipe lines have been functioning from 26 to 34 years, and through the use of special cleaning tools which were developed at the plant, as well as deaeration of the more active Toconce water, these pipes are now maintained at capacities which do not diminish as years go on. Constructing the Dam The Arroyo Salado pipe line design and construction involved certain special and interesting features, and inasmuch as this line and its intake works are solely for the needs of the new sulphide plant, more detailed description is given. The waters are impounded at a gravity dam constructed of concrete to a height of 100 ft above the river bed, keyed into the precipitous Dacite walls of the narrow canyon (barely 6-ft wide at the bottom, only 25-ft width at 50 ft above). A small secondary dam was built 100 ft down stream from the main dam, providing a pool of 15-ft depth to protect the main structure from flood flows over the spill-way during the rainy season. A system of four 36-in. syphons was designed for discharging these flood waters from the lower depths of the lake, in order to avoid eventual sedimentation behind the dam. The lake has a length of 3300 ft, and its water level is controlled by an adjustable spillway permitting draw-down of eighty inches, amounting to 41,000 metric tons of available capacity. This regulation is necessary because of wide fluctuations in stream flow between day and night due to freezing of feeders. During the construction of the dam the entire river flow was handled within a 36-in. pipe line some 2000 ft in length. As the excavations proceeded

Jan 1, 1952

By H. P. Rassbach, E. R. Saunders

MUCH progress has been made in recent years in the theory and practice of making stainless steel. By effective utilization of oxygen for decar-burization and more suitable alloying agents, it has been possible to attain consistent production of very low-carbon stainless steel. In order to facilitate economical production of stainless steels, the Electro Metallurgical Co. has carried out an extended experimental program that has clarified some of the complex interrelations of temperature and composition under decarburizing and reducing conditions. These results'-:' have been founded in large part on small experimental heats, and in order to confirm their validity and significance under commercial conditions, a survey of stainless steel melting practice has been made with the cooperation of stainless steel producers. Conditions required for decarburization and associated oxidation of chromium, manganese, and iron have been fairly well established in relation to the beneficial effect of the highest practicable temperature. The recovery of chromium and manganese from highly oxidized slags by reduction with silicon has been indicated with somewhat less precision and this study has shown significant deviation in large commercial heats from the small experimental heats. In spite of an unsatisfactory degree of accuracy in estimating the conditions affecting reduction of metallic values, it has been possible to calculate a slag weight in reasonably good agreement with the results observed in large heats. While there still is much to be learned about both the qualitative and quantitative aspects of stainless steel melting, this survey has indicated the manner by which the efficiency of the production process may be measured. Oxidation Period In order to establish practices for the recovery of chromium and manganese from the slag the amounts of these metals oxidized should be known. Moreover, since reduction of oxidized chromium and manganese must necessarily be accompanied by similar reduction of iron, knowledge of the total quantity of metallics oxidized is essential. The relations between carbon, chromium, and temperature under oxidizing conditions were developed by Hilty in 1948.' While it had been realized previously that retention of chromium in the metal during carbon oxidation is favored by high temperatures, the Hilty relation provided quantitative information useful for evaluating melting procedures. For example, it was shown that decarburization to moderate carbon levels can be achieved while retaining substantial amounts of chromium. However, in decarburizing to the low-carbon level necessary for producing 0.03 pct maximum carbon stainless steel, very little chromium remains in the bath at the temperatures generally employed. For this reason, Hilty, Healy, and Crafts' later projected an extension of the chromium-carbon relation to the low-chromium region. Although this chromium-carbon-temperature relation defined the composition of a chromium steel bath at the end of the oxidizing period, it did not provide a direct means for estimating the total amount of metallic oxidation occurring during decarburization of a stainless steel heat. This subject was investigated by Crafts and Rassbach³ and, later, by Hilty, Healy, and Crafts" who demonstrated that metallic oxidation is a function of the chromium content of the initial furnace charge, the minimum carbon content attained, and the temperature. It was further demonstrated from data on 1-ton heats that an empirical relationship exists between the ratio of chromium plus manganese to iron in the slag (S) to the corresponding ratio of these components present in the metal bath. The following expression was derived to permit the calculation of the amount of metallics oxidized during decarburization of a given charge: W-2000 (Cr1 + Mn1)-(Cr2 + Mn2) 1 100s/s+1-(Cr3 + Mn2) where W is the pounds of chromium, manganese, and iron oxidized per ton of charge; Cr1, the percentage of chromium in the charge; Mn1, the percentage of manganese in the charge; Cr2, the percentage of chromium in the bath after oxidation; Mn2, the percentage of manganese in the bath after % Cr + % Mn oxidation; and S =%cr+ % Mn/%Fe in the slag after %Fe oxidation. In order to establish whether the slag ratios of commercial heats are consistent with those found in the experimental heats, the slag and metal analyses shown in Table I were evaluated. These data represent samples taken after the oxidation of two 1-ton and seven 15 to 25-ton heats of types 303, 304, and 304L stainless steel made in chromite, acid, and basic hearths. Fig. 1 illustrates that the results for the commercial heats correlate reasonably well with the relationship originally established for experimental 1-ton heats. It will be noted that in the range of metal composition of these heats, the slag values generally lie above the line. As was suggested by

Jan 1, 1954

By W. H. Smith

IN connection with some recent work on the effect of impurities on the ductility of chromium, it appeared desirable to know the solid solubility of carbon in chromium. A literature survey indicated that this information was not available. Although considerable work has been done on the Cr-C phase diagram,&apos;.&apos; previous investigators have been more concerned with the structure and phase boundaries of the carbide phases than with the terminal solid solution. The phase diagram shown in Fig. 1 is taken from the work of Bloom and Grant&apos; and represents the most recent determination. As indicated by the dashed line, the a solid-solubility limit was not determined. Experimental Procedure Alloys of chromium were prepared from hydrogen-treated and vacuum-degassed electrolytic chromium plus spectrographic grade carbon. The oxygen and nitrogen content of the alloys was <0.002 pct. After melting, analysis of the alloys showed them to contain 0.02, 0.08, 0.15, and 0.55 pct C. Pieces of the alloys were heated in a protective atmosphere to various temperatures and then quenched after holding for a time sufficient to insure equilibrium. Microscopic examination of the as-quenched alloys for the presence of a second phase was used as a measure of the solubility limit. The heats were made in a multiple hearth arc-furnace using a zirconium-gettered static argon atmosphere. A zirconium melt was made before each Cr-C heat. Triple melting was used to insure ingot homogeneity as shown by microscopic examination. The alloys were prepared by adding portions of a 4.5 pct C-Cr master alloy to high purity chromium. The carbon contents listed previously were those obtained by analysis. The nitrogen and oxygen contents after arc melting were both <0.002 pct. Sections 1/8X1/4X1/2 in. were cut from the 100-g ingots and a hole drilled in one end in order to suspend the sample from a molybdenum wire. After the surface was carefully cleaned, a sample of each melt was hung in a mullite tube heated externally by a platinum resistance furnace connected to a vacuum system. The lower portion of the mullite tube was sealed to Pyrex and closed off several inches below the furnace. This was filled with sili-cone oil kept cold by circulating cold water around the outside of the Pyrex. Quenching into the oil bath was achieved by melting a fuse wire supporting the sample. It required about 4 sec for the sample to cool from 1400" to 600°C. This severity of quench was considered satisfactory to freeze-in the high temperature equilibrium. For tests made at temperatures of 900" to 1200°C, heating was done in vacuum; for tests above 1200°C, an argon atmosphere was used. The holding time employed ranged from 12 hr at 900°C to 6 hr at 1400°C. Experiments were performed at temperatures of 900°, 1000°, 1100°, 1200°, 1300°, and 1400°C. Microscopic examination for evidence of a second phase was done at X1500. Experimental Result The microstructures of a 0.08 pct C-Cr alloy as-cast and after quenching from 1300°C are shown in Figs. 2 and 3. A 0.15 pct C-Cr alloy quenched from 1300°C is shown in Fig. 4. The data obtained from the quenching experiments is shown graphically in Fig. 5. If the Van&apos;t Hoff equation is obeyed, a plot on a logarithmic scale of the mol fraction of solute vs the reciprocal of the absolute temperature should give a straight line. For dilute solutions the weight percentage can be substituted for the mol fraction without introducing any appreciable error. The Van&apos;t Hoff equation can then be written as where H is the heat of solution in calories per mol. The slope of the straight line on the log pct C vs 1/T plot gives the value of AH. Assuming that the Van&apos;t Hoff equation is obeyed, which is probably justified for the dilute solution of carbon in chromium, the heavy straight line shown on Fig. 5 represents the best fit of the data. This line was obtained as follows. On Fig. 5 the results of the microscopic examination of all alloys following quenching were plotted and designated as to whether one or two phases were seen. Below 1100°C all alloys showed a second phase on quenching. The heavy vertical lines shown in Fig. 5 therefore represent the possible range of the ter-

Jan 1, 1958

By C. W. Thompson

Paul D. Suloff (Goodyear Tire and Rubber Co., Inc., Akron, Ohio)—I would like first to comment on problems of the conveyor belt discussed in Mr. Thompson&apos;s excellent paper, since that is what we hope we know most about. Twists in relatively wide conveyor belt unavoidably produce a lateral maldistribution of tension, raising tension at belt edges and reducing it at the center. They also produce a lateral collapsing force on the belt at the center of the twist owing to the inherent tendency of all the longitudinal elements of the belt to try to pass through a point at the twist center. Calculation of the twist geometry by the methods shown in Mr. Thompson&apos;s paper keeps these extraordinary forces within limits which the belt designer can tolerate. No reduction in belt life due to twisting need be contemplated when this geometry is maintained. There is a minor exception that belts of extreme lateral flexibility will tend to curl laterally at the center of the twist. However, any ordinary fabric construction will perform satisfactorily in this respect. These twists are always made in regions of low tension in the conveyor so that even in the edges of the twist, belt tension does not exceed the average tension found in highly stressed regions of the conveyor. Offsetting these out-of-ordinary belt stresses is the advantage that Mr. Thompson has brought out of getting the return run up out of the dirt where it can be seen. This not only makes it easier to train, but also, in the event that it is not properly trained, frees it of the normal return run edge wear hazard. It is well known that return run edge wear is a prominent cause of belt mortality underground. An interesting aspect of this two-way conveyor is that the belt may be made what is known as a Mobius Strip. A Mobius Strip is obtained by splicing a belt after turning one end of it 180" about its longitudinal axis. In other words, one end is turned upside down before splicing. A belt spliced in this fashion turns itself upside down every time it comes around, but the twist which has been put in the splicing, of course, stays at one location on the conveyor, in this case one of the twist sections at the end. Turning the belt over every revolution might have advantages in some cases. Belts could be made with equal covers and the two sides worn equally and simultaneously. In this case there would be no problem of getting belts on upside down by mistake. However, the two-way conveyor does not have to be a Mobius Strip. It can be twisted in such fashion that the same side is up on both runs. It is simply a question of which way the final 90" twist is made before joining the ends. Another interesting aspect of the two-way conveyor is the problem of operating two-way conveyors in series. Here the sequencing of starting brings up some new problems. It will be recognized, although not always at first glance, that if the starting sequence is planned for one run of the conveyor the reverse will result on the other run. With the two runs carrying bulk material in both directions a reverse sequence on one run would be intolerable. In this situation the only solution appears to be a simultaneous starting of all conveyors in the series. However, with the coal in one direction and intermittent supplies in the other it would be entirely practical to sequence the conveyors for the coal run and accept a reverse sequence on the supply run. The two-way conveyor also lends itself to new driving possibilities. First, it is quite possible to drive at the head end of each run, which of course, means a drive at each end of the two-way conveyor. Driving in this way a given belt can be extended to substantially greater lengths than a conventional conveyor with drive at one end only. In addition to this, under certain conditions the conveyor can be extended to extreme length by driving at one end and at some intermediate point on the most heavily loaded run. As a particular case, a belt carrying coal downgrade and supplies back upgrade could be extended to extreme lengths by driving at the head of the coal run and at an intermediate point of the supply run. Mr. Thompson has been a pioneer in belt conveyor transportation underground and his accomplishment here with the first two-way conveyor of any consequence is another notable addition to the art.

Jan 1, 1954