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By F. R. McDonald

The proton NMR spectra of 32 pyridines in two solvents, carbon tetrachloride and benzene-d6, are presented. The chemical-shift and differential-shift values are correlated with the substitutent position and with the solvent media. The ring protons and the protons on alkyl substituents are tabulated. The paramagnetic shift of the protons adjacent to the nitrogen is discussed, and a proposed structure for the pyridine-benzene complex that is consistent with the observed anomalous paramagnetic shift is presented. The utility of the data in the identification of pyridine compounds found in shale-oil naphtha is demonstrated.

Jan 1, 1971

By Keenan. Albert M

The challenge of mining coal seams having a pitch of 30 degrees is one not frequently found in the United States, due primarily to the scarcity of such deposits. However, in the Thompson Creek District in Pitkin, Colorado, such a challenge exists to produce high quality medium volatile coking coal, and development of a longwall system is underway. The deposits under consideration are those of the Anschutz Corporation Inc. of Denver, Colorado, which are not in a virgin state. The properties were originally mined by the Colorado Fuel and Iron Corporation from 1894-1907, and by the Thompson Creek Coal and Coke Corp. from 1953-1966. Approximately five and one half million tons of coal have been mined by the above companies, attesting to its quality. The cessation of operations in both instances was occasioned by economics. Both companies employed a room and pillar system of mining, but being denied the use of high capacity mining and loading machinery because of the excessive pitch of the coal seam. This resulted in an unfavorable productivity expressed as output per man-shift, and thus the high cost of production occasioned by hand loading and semi-mechanization dictated cessation of operations. The specific location of these deposits are in Pitkin County,

Jan 1, 1971

By D. G. Kesterke

Research was conducted to determine the feasibility of extracting elemental oxygen from silicate-bearing melts by high-temperature electrolysis. The method was based on a modification of existing processes for electrowinning metals from their oxides dissolved in molten halide systems. Electrolyses were performed at temperatures ranging from 1,050° to 1,250° C in boron nitride cells that incorporated a silicon carbide cathode and an iridium anode. The electrolytes contained volcanic scoria as the oxygen source, plus mixtures of BaF2 and LiF as fluxing agents. Cell gases containing almost 14 volume-percent oxygen were obtained. The average anode current efficiency for several experiments of extended duration was calculated to be 55 percent. Cell design and operating conditions are described, and some of the problems associated with the process are discussed.

Jan 1, 1971

Samples weighing 0·5 g are fused with potassium hydrogen sulphate, leached with 4 ml. of 4 M nitric acid and diluted to 10 ml. with water. An aliquot of the clear solution is treated with EDTA, adjusted with ammonia to give a pH value between 5 and 6 treated with cyanide, ammonia, and sodium diethyldithiocarbamate then bismuth diethyldithiocarbamate is extracted with chloroform and the yellow colour is compared visually with a standard series. Alternatively, the bismuth complex may be filtered on a "Millipore" filter and subjected to x-ray fluorescence spectrometry.

Jan 1, 1971

By P. G. Pigott

This Bureau of Mines laboratory research demonstrated that building bricks meeting ASTM specification C62-66 for grades SW and MW of FBS-type bricks can be produced from waste copper mill tailings. Waste tailings from accumulations at Magna, Utah, Anaconda, Mont., and Hayden, Miami, and San Manuel, Ariz., were dry-pressed into bricks, and the products were tested in accordance with ASTM Designation C67-66 on sampling, and testing of brick. An evaluation derived from pilot-plant production of similar brick at the Coal Research Bureau of West Virginia University indicated the production cost to be about $34 per 1,000 bricks. Utilization of such mill tailings on a commercial scale would eliminate the need for digging new pits for brick raw material production, conserve natural resources, and simultaneously remove potential sources of air pollution.

Jan 1, 1971

By George A. Grandy, D. W. Fuerstenau

Simulation of nonlinear grinding systems is discussed in the context of the size-discretized batch-grinding model. A linear approximation of environment-dependent (or nonlinear) selection functions is postulated for systems which exhibit feed grinding orders less than unity. For this, a convex linear combination of zero and first-order selection function forms is used to obtain an analytical solution that can be used for simulation. An example in which marked nonlinear behavior occurs is rod-mill grinding. Accordingly, a series of grinding experiments was carried out in a batch laboratory rod mill to test these techniques; agreement between experiment and the model was found to be very good. Recently, significant progress has been made in the simulation of linear grinding systems by using a population balance mathematical model. Herbst and Fuer-stenaul presented a parameter estimation scheme that can be used with the size-discretized batch-grinding model, and Gumtz and Fuerstenau3 have extended these techniques to the simulation of locked-cycle grinding. Grandy et aL." summarized the simulation of linear systems and noted important alternate solutions to the batch-grinding equation. Kelsall et a1.'-" discussed the dynamic simulation of linear batch grinding, as well as linear open and closed-circuit grinding. In all of these, the key word is linear, which certainly restricts the engineer trying to simulate a variety of grinding systems. This paper discusses an approach for the simulation of non-linear grinding systems, particularly those which exhibit grinding orders less than unity. Experimentally, dry rod-mill grinding is used to evaluate the technique presented. Batch Grinding Model The size-discretized batch-grinding model has been discussed in detail by Mika et al.'.' and by Herbst and Fuerstenau,' and therefore, will only be briefly developed here. The model is formulated by making a differential mass balance of each size fraction in such a manner that the equation parameters are precisely defined in terms of selection and breakage functions,' which can be either calculated'.' or estimated1 from experimental data. The model is phenomenological because it averages all stress-application events in the system and, therefore, does not distinguish the manner in which individual particles are stressed.' For this average environment, the resistance of all particles in a size fraction to breakage is defined by that size-frac-tion's selection function; and when fracture occurs from an average event, the breakage product size distribution is defined again by that size-fraction's breakage function. The model is deterministic and is derived in terms of a continuous time variable and a discretized size variable,' i.e., the n mass fractions associated with the n defined size intervals (see Table 1 for nomenclature).

Jan 1, 1971

By M. Gyenge, O. F. Coates, K. Barron

"It is suggested that artificial supports might be used to increase, with safety, the angle of hardrock open-pit slopes. Guidelines for designing such a support system, based on a simplified analysis, are given. The installation, construction and instrumentation of a trial support system is described and a breakdown of the construction costs is given. On the basis of this cost analysis, the economic benefits derived from reduced waste-rock excavation are compared with the support costs for a typical mine. In an example, it is shown that the slope angle in an open-pit mine 500 ft deep, with benches 50 ft high by 30 ft wide, could be safely increased from 37Yz degrees without support to an optimum of 53 degrees with support. In this case, the support costs would be approximately $900 per linear foot of pit wall and the savings due to decreased excavation of waste rock (at $0.34/ton) would be approximately $1,900 per linear foot, yielding an increased margin of profit of approximately $1,000 per linear foot of pit wall. This figure could be increased to $6,000 per linear foot if the steeper slope were used to extract more ore.More than 50 per cent of the support costs are accounted for by diamond-drilling the holes for the deep cable anchors. The use of percussion drills for this purpose would seem to offer a means of significantly reducing the support . costs and thus further increasing the profit margin available."

Jan 1, 1971

By Peter N. Calder

The selection of a slope angle for an open pit in rock has often been made without the benefit of a comprehensive engineering study. This is largely due to the fact that the hardware and procedures necessary for such a study have not been available. Recently this situation has been reversed. The application of borehole television and terrestrial photogrammetric devices has made possible the accurate determination of rock joint structure, on which stability calculations can be based. Another important aspect is the evolution of practical methods of conducting in-situ rock joint strength tests.

Jan 1, 1971

By R. R. Montigny

A brief description is given of the history, location and geology of the Granisle Copper Limited operation on Copper I land in Babine Lake. Detailed information is given with regard to pit design, planning, drilling, blasting, excavation and haulage. Equipment maintenance and mining costs are also presented.

Jan 1, 1971

In his introductory remarks Mr. MacKellar referred to three types of development which may be considered to compete with mining for land use. 1. Total Development - in which he cited urban development, and intensive agriculture. 2. Partial development - in which he cited pastoral, and in this category would come National Parks. 3. Non development - in which category would fall the primitive or wilderness areas.

Jan 1, 1971

By Ralph B. Adair

The Bureau of Mines conducted laboratory batch and small-scale continuous tests on samples of silt from the Nolichucky Reservoir near Greeneville, Tenn., to develop a process for the recovery of commercial-grade muscovite concentrates. Using a combination of Humphreys spirals for preconcentration followed by cationic flotation, 94.5 percent of the muscovite was recovered in a product containing 96.5 percent muscovite.

Jan 1, 1971

By Richard Chojnacki

The western united States is not usually noted for the production of fluorspar; however, many significant fluorspar districts do occur in the Rocky Mountain region and constitute a resource of sizable dimension and major potential. This is particularly true within the area of the West served by the rail lines of the Union Pacific Railroad (see Figure 1). Fluorspar deposits near the Union Pacific have produced over 1.3 million tons of finished fluorspar products from 2.3 million tons of ore. Approximately 80% of this production was of acid-grade and 20% of metallurgical-grade fluorspar. Production of ceramic-grade was negligible. Through the years substantial production has come from eight mining districts. Five of these have produced metallurgical-grade and three acid-grade. The best-known fluorspar producing area in the West is the Jamestown Mining District. It was one of the first producers of fluorspar in the western United States and is estimated to have produced over 500,000 tons of finished fluorspar from over a million tons of ore. The district encompasses approximately 36 square miles in central Boulder County, Colorado, about 9 miles northwest of the City of Boulder. Although small shipments were made from the district as early as 1873 or 1874, significant sustained production was not achieved until 1940 when the General Chemical Company, now a division of the Allied Chemical company, acquired the Alice, Burlington, Chancellor,

Jan 1, 1971

By R. L. Rough

Photoelastic, wire resistance, and borehole-deformation gages were used to measure the in situ strains when an outcropping Black Hand Sandstone Formation in Hocking State Forest, Hocking County, Ohio, was overcored to stress-relieve the surrounding sandstone. Strains transmitted to the photo-elastic gage did not produce isochromatic patterns suitable for quantitative analysis. Strain readings from the wire resistance gage indicated tensile-strain loading during overcoring; readings from the strain-gage bridges in the borehole-deformation gage indicated compressive-strain relaxation during overcoring. The magnitude and orientation of the computed stress fields obtained from both gages were comparable when the corresponding tensile and compressive elastic moduli were applied. Stresses computed from deformation and resistance gages at depths to 5 feet showed the maximum stresses to range from 1.042 to 2.121 times greater than the minimum stresses. The bearings of principal stresses calculated from the deformation gage readings varied between N 61° E and N 83° E; the wire resistance strain gage bearing computations varied between N 68° E and N 82° E. Measurements of in situ strain in sandstone were shown to have promise but to require further development in field technique that could likely be achieved through experience.

Jan 1, 1971

By Robert V. Whitman

INTRODUCTION Ideally, this paper should start by citing examples of slope failures in open pit mines during earthquakes. I do not know of any such examples, although - since my search of the literature has not been exhaustive - such examples may well exist. However, there certainly are many examples of devastating earthquake-induced landslides, and there is no reason why such landslides cannot occur at large open pits. Moreover, many of the actual earthquake-induced landslides have seriously affected mining operations in one way or another. Hence, this paper will treat earthquake-induced landslide phenomena in general, in the hope and expectation that the reader can apply this general knowledge to the specific problems associated with open pit mining. Ultimately, of course, the aim is to predict the likelihood of the occurrence of earthquake-induced landslides, and to delineate steps which can be taken to minimize the possibility of their occurrence. These aims can be achieved only imperfectly at the present time, for our knowledge concerning earthquake-induced landslides is still in an embryonic stage. Nonetheless, much useful knowledge does exist. In order to understand and apply this knowledge, it is necessary to: 1. Review the nature and extent of actual earthquake-induced landslides; 2. Have a grasp of the basic terminology and concepts of seismology; 3. Study the fundamentals of probability-based seismic risk analysis; 4. Appreciate the way in which earth materials behave during earthquake loadings; and 5. Have a general knowledge of the methods currently available for analyzing the dynamic response of slopes. Much of this paper is devoted to this necessary background. Information of practical use in connection with various classes of slope stability problems is assembled in the final sections. LANDSLIDES DURING PAST EARTHQUAKES Almost every earthquake in mountainous country, strong enough to be reported in the literature, has caused at least some landslides and rockfalls. Unless these slides or falls damage

Jan 1, 1971

In recent times there has been a marked increase in public awareness of air pollution. Some of this has been induced by the Press but our own senses tell us that there is a visible haze over our cities on many mornings. The people are now demanding a clean environment and will no longer tolerate smoke and dust fall or gaseous and odorous emission which either soil building and furnishings or interfere with comfort and well being.

Jan 1, 1971

By Y. Okahara, I. Iwasaki

Chlorination behaviors of pure iron and manganese oxides were investigated by combining a thermogravimetric analysis (TGA) technique with batch-boat roasting followed by leaching. Ferrous and manganous oxides could be chlorinated readily, but, in the absence of a reductant, the higher oxides of both iron and manganese were dificult to chlorinate. Thermogravimetric analysis curves were drawn to illustrate the complexities of the reactions, and the possible mechanisms were discussed. Then three manganif-erous materials from the Cuyuna Range of Minnesota were treated by a process involving the selective chlorination of manganese followed by leaching. The results were interpreted in the light of the chlorination mechanisms observed on the pure iron and manganese oxides. One of the unsolved problems challenging Minnesota's mining industry is the development of an economic process to utilize the manganiferous iron ores of the Cuyuna Range. Although these ores constitute one of the major domestic reserves of manganese (estimated at 432 million tons'), their grade is too low for them to be mined solely for manganese. The nearly exclusive dependence of the U.S. steelmaking industry on foreign manganese ores has stimulated research over the years in treating the low-grade Cuyuna ores. As a result, a number of processes have been proposed.'-' Among these processes is chlorination. Chlorination is the high-temperature conversion of metallic oxides to chlorides by either chlorine or hydrogen chloride gas, or by solid chlorides of sodium, calcium, or magnesium. The metallic chlorides are then recovered by either vaporization or leaching. Although chlorination is an old art, it has become popular in recent years for the recovery of such new metals as titanium, zirconium, and niobium; for the removal of impurities like copper, lead, and zinc from pyrite cinder for ironmaking raw materials;' and for the segregation roasting of coppera and nickel' ores. The chlorination processes proposed for manganifer-ous iron ores have been relatively few (hydrogen chloride chloridization-leaching, hydrogen chloride and calcium chloride chloridization-volatilization,' and chlo-rination-selective condensation6), due presumably to the economics and to the corrosion problems involved. However, the successful application of chlorination processes to other ores will undoubtedly contribute to technological advancements in the chlorination of man-

Jan 1, 1971

By R. F. Waters

The Bureau of Mines developed laboratory-scale processes to recover sodium phosphates and metal values from waste sludges generated by the phosphate treatment of automobile and appliance components. After treatment of the sludge with NaOH, followed by coke or hydrogen reduction and sintering, zinc and lead were recovered as fume; the sinter was separated by leaching into a metallic product, chiefly iron and nickel, and high-purity disodium or trisodium phosphate.

Jan 1, 1971

By Robert S. Shoemaker

The development of a metallurgical project is a business venture rather than a technical exercise, and its real objective is to obtain the maximum return on investment consistent with the limitations imposed by company policy and the requirements of society. Demand for profits is the driving force for technical advances and naturally leads to improvements in processes, equipment and plant designs. In the mineral dressing field there have been virtually no new processes developed since the introduction of flotation, and the steady erosion of profits by inflation has been offset only by increasing equipment size. Unfortunately, in recent years, plant designers have not always utilized the largest equipment or the opportunities for functional and economic design presented by the development of larger equipment. Lower cost plants can be developed providing that unrealistic restraints arising from the prejudice of past practices are avoided. As an illustration of this premise, three different approaches were taken toward the design of a clearly delineated but hypothetical copper concentrator of 20, 000 tons per day capacity. While the same flowsheet was used in each case, designs were produced for a conventional plant, an alternate using the largest available equipment, and a third scheme utilizing large equipment and the most functional design which could be developed, including features derived from many plants in many different industries. Capital costs were developed for the three separate designs based on a location in the southwestern United States. Ancillary facilities were identical in each case and included a mine truck shop, a maintenance shop, a general office, and a nominal starter dam for tailings disposal within 5000 feet of the plant. Water supply within 1000 feet of the plant, and power furnished to the property boundary has been assumed in each case. The flowsheet of the three plants is shown in Figure 1 and represents the treatment of a typical sulfide copper ore by three- stage crushing followed by single-stage ball milling. Concentration is by flotation roughing and scavenging followed by regrinding of the concentrates and two stages of cleaning.

Jan 1, 1971

By H. A. Dierks

The work and costs involved in controlling three large fires in abandoned mines in the anthracite region of Pennsylvania are presented in detail. The work was done under authority of Public Law 89-4. Three basic control techniques were successfully employed by the Bureau of Mines: (1) forming underground fire-containing barriers by systematically flushing incombustible material into mine voids, (2) excavating an isolation trench around the fire area, and (3) complete excavation of hot and burning material in the fire area. When methods 2 and 3 were used, the excavations were subsequently back- filled with incombustible material. Preliminary location of fire perimeters by the standard use of patterned 6- and 8-inch-diameter boreholes is discussed. Investigations into air-borne infrared scanning, shallow subsurface thermal anomalies, and the use of helium to trace the movement of mine air currents are covered. Although they are costly and time consuming, boreholes have proven to be the most efficient means to date of obtaining necessary subsurface fire data. Infrared scanning was ineffective in delineating fires under more than 30 feet of cover. Helium-tracing was shown to have a good potential for determining possible hazards to adjacent residential areas from traveling mine fire gases.

Jan 1, 1971

A digital computer program has been developed to mathematically simulate the effect of limited changes in feed rate on the grinding circuit of an operating mill. The data on which the program is based was supplied by the staff of the mill in question. Samples taken around the complete grinding circuit were analyzed for solids content and particle size distribution for two feed rates. Matrix models were used to represent the grinding units: rod mill and pebble mills. Parameters for these models were computed using a method developed by Mular(l). Classifier parameters were computed by the Powell search method(2) for the Plitt model (3) . The simulation program, which permits the computing of the size distributions in the circuit resulting from the changes in feed rate between the two measured values, assumes that a linear relationship exists between the corresponding parameter coefficients. The program uses an iteration method to bring the simulated grinding circuit to a steady-state material balance for any specified' feed rate between those measured. The program is, in addition, extremely fast and economic in computing time. The results show that the simulations are not everywhere consistent with practice. For example, higher feed rates on simulation predict finer products. These errors are shown to be attributable to anomalies in data gathering techniques. An improved sampling procedure for grinding-circuit data gathering is suggested for future simulations of this type. It is concluded that the

Jan 1, 1971