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By J .. R. Dalton

"Blasting in high sulphide content ores presents at least two major safety considerations: (1) compatibility of the explosives with the ore while resident in the borehole; and (2) introduction of a flame and shock producing explosive into this environment, while minimizing the risk of reactive ignition in the chemical reaction products/sulphide dust cloud resulting from the detonation. This paper will discuss only the second consideration, i.e. post detonation reactivity.The ETI (Explosives Technologies International) Technical organization, in consultation with the U.S. Bureau of Mines, reasoned that the most relevant considerations in the development of an explosive for this application are:• minimization of reactive gaseous detonation reaction products;• flame duration reduction;• flame temperature reduction;• increased explosive sensitivity;• reduced probability of ""low order"" detonation; and• maintenance of energy level as high as possible.Obviously, not all of these are totally independent properties, i.e. increased sensitivity will reduce the probability oj low order detonation as will higher energy levels. However, these considerations were examined separately as if they were independent."

Jan 1, 1994

By Raymond M. Coveney

Bedded sulfide ores in Cambrian Chinese black shales typically contains at least 2% Ni, 4.5% Mo, 1% As and 0.2% Se. Thin laminae and replaced on colites from the ore contain intimate mixtures of fine-grained Ni-rich sulfides and jordisite complicating ore treatment. The ores contain abundant jordisite, vaesite, nickelian pyrite, and barite and lesser amounts of sphalerite, gersdorffite, millerite, tennantite, fluorite, chalcopyrite, and HgSe (probably tiemannite). At present only ores that are sufficiently rich (>4.5% Mo) for direct roasting are mined and treated to produce, ammonium molybdate. The ores contain Pt, Pd, and Au in amounts totaling up to -2.5 ppm, but these are not recovered.

Jan 1, 1993

The original work statement of this contract was entitles: 'A Accident Analysis Model for Use in Underground Coal Mines.' The contract was scheduled for a duration of 26 months and was to be in three phases: Phase I - Model Development, Phase II - Application of the Model, and Phase III - Installation, Monitoring and Evaluation of the Model. Alter three months in Phase I, it became apparent that much of the needed data for the model was nut available and could not be made available within the original study schedule. Proceeding as plumed would have resulted in a model of limited utility. Personnel from Theodore Barry and Associates and the Bureau discussed the problem in detail and mutually decided to utilize the majority of remaining funds to assist the Bureau in solving these data problems and develop a stronger ongoing capability for safety analysis within the Bureau. Both tie contractor and the bureau officials involved believed that this decision greatly enhanced the Bureau's ability to develop future accident prediction models, whose rigor would greatly exceed what was currently possible.

Jan 1, 1972

By R. S. Silo

A new process for leaching oxidized copper ore (CuO) is presented. Its principle advantage is that it requires low cost raw materials. For large operations the only materials used are sodium chloride (NaCl) and calcium carbonate (CaCO3). In those operations of medium size, sulfur, which is abundant in the North of Chile, is used. However the process may also be applied to small copper plants provided low cost sulfuric acid is available as a substitute for sulfur. That is, it may be applied to those small capacity plants where large additional installations for the regeneration of the leaching solution are not needed. The method is comprised of six stages, namely: Crushing, Leaching, Washing of the Solution, Precipitation, Regeneration of the Leaching Solution, and Fusion. A flow diagram for this process in a large or medium-sized plant is given. LEACHING This process requires hydrochloric acid as the leaching solution for treating copper ores. This is easily produced from sulfuric acid and sodium chloride as follows: A small excess of NaCl is advisable to avoid the presence of free sulfuric acid. The experiments reported here were performed with 36 grams of HC1 per liter of solution to obtain a free acidity equivalent to 50 gpl of H2SO4 (the usual leaching solution concentration). Hydrochloric acid leaching may be done by either of the usual methods; i.e. percolation or agitation. The method selected depends on the characteristics of the ore. No matter which method is chosen, low pressure air can be used either as an air lift for circulation of the solution, or for agitation of the pulp with "Pachuca" tanks. In either case, there are no problems with respect to oxidation. Leaching by agitation produces good recovery; but at times, the problem of high consumption of leaching solution or the overproduction of grinding slimes which interfere with contact between phases negates this advantage. The percolation method was used in this study. Percolation took place in four 15-liter containers, 11 1/2 in. diam. by 12 in. height, having hemispheric bottoms. Each container was fitted with a wooden bottom having 1/4 in- holes. The air lift was a 3/4 in. plastic tube 10 in. high. Air at 6 psi was supplied from a rotating laboratory compressor. 10 kg of ore were charged to the container and counter-current leaching was used. The leaching solutions were fed through the center pipe in order to use the ascending solutions to reject the air in the charge. The solution circulated through the air lift in order to facilitate the movement according to the increase in density produced by the reaction between the phases present. Specific observations made in this report cannot be generalized because the leaching characteristics are dependent on the nature of the ore used, the method of leaching, and the production of primary or secondary slimes while crushing or leaching.

Jan 1, 1967

By J. P. Manning

UNQUESTIONABLY, the outstanding feature of the piping for the sulphide plant is the large amount which had to be done in almost every size from instrument tubing to 84 in. OD pipe. In this article the piping will be considered under the main headings of Yard Piping and Building Piping. Tables I and II give data concerning the individual systems. In the case of yard piping the figures pertaining to lengths and weights are approximately correct; those for building piping are given only for the concentrator and smelter buildings where the amounts were the most extensive. Yard Piping The installation specifications for yard piping called for all sizes larger than 12 in. to have welded joints made with outside welding rings. This type of joint was selected over butt-welded joints because less time would be needed for lining up, damaged pipe ends would be less of a problem, and deflections up to 6° per joint could be made when necessary. These advantages were thought to outweigh the disadvantage of the greater amount of welding required. Yard piping in sizes of 3 to 12 in. was specified to have butt-welded joints made with inside welding rings which were intended to facilitate lining up and keep welding material out of the lines. Screwed connections were specified for sizes under 3 in. No protective coatings or wrappings were required on pipe to be buried because of the dry climate at Chuquicamata.

Jan 1, 1952

By H. Conrad, V. V. Damiano, G. J. London

A method is described for producing single crystals of high-purity beryllium, Be-4.37pct Cu, and Be-5.24 pct Ni. These crystals were prepared for testing in compression parallel to the [0001] by orienting and lapping to within ±3' of arc of the (0001). Microstrain testing apparatus is described along with c axis compression results for ingot purity beryllium, twelve-zone-pass material, and the above-mentioned alloys. Results show no measurable plasticity for the ingot purity material from -196" to 400°C, although some surface traces of (1122) slip was observed at 200°C and above. The twelve-zone-pass material shows substantial microstrain plasticity at 220°C with slip on (1122). Both alloys show significant plasticity at room temperature and above with slip also on (1122) planes. THE two slip systems which normally operate during the plastic deformation of beryllium in the vicinity of room temperature are:' basal slip (0001)(1120) and prism slip . Pyramidal slip with a vector inclined to the basal plane has been reported for elevated temperatures,'-a but occurs near room temperature only at very high stresses.~ A summary of the available data on the effect of temperature on the critical resolved shear stress for slip on these systems has been compiled by Conrad and Perlmutter.~ It has been postulated6'7 that one of the principal factors contributing to the brittleness of poly crystalline beryllium at temperatures below about 200°C is the difficulty of operating pyramidal slip with a vector inclined to the basal plane. Hence, detailed information on the operation of such a slip system is important to understanding the brittleness of beryllium. The operation of pyramidal slip with a vector inclined to the basal plane is best accomplished in beryllium by compressing single crystals in a direction parallel to the c axis. In such a test the resolved macroscopic shear strzss on the basal and prism planes is zero and (1012) twinning which is favored by tension along the c axis does not occur. Hence, in c axis compression of beryllium the normal deformation modes are inhibited and the operation of pyramidal slip with a vector inclined to the basal plane is favored. In the present investigation, c axis compression tests were performed on beryllium single crystal as a function of temperature (77" to 700°K), purity (commercial and twelve zone pass), and alloy content (4.37 wt pct Cu and 5.24 wt pct Ni). Presented here is a description of the test techniques employed and the gross mechanical behavior observed. A detailed analysis of the slip traces developed on the surfaces of the deformed specimens during these tests and the results of electron transmission studies of the deformed crystals are given in a separate paper.B PROCEDURE 1) Materials and Preparation. Single crystals about 1 in. diam were prepared of the following materials: commercial-purity beryllium, high-purity beryllium, and two beryllium alloys, one with 4.37 wt pct Cu and the other with 5.24 wt pct Ni. The commercial-purity single crystals were obtained by cutting specimens from large-grained ingot of Pechiney SR material, which is approximately 99.98 pct pure. The high-purity crystals were prepared by floating-zone refining (twelve passes) a rod (7 in. by 1 in, diam) of Pechiney SR grade cast and extruded beryllium. Although an absolute chemical analysis of the zone-refined material was not established, mass spectro-graphic analysis, emission spectrographic analysis, and y activation analysis indicated that it contained in atomic fractions about 5 to 10 ppm each of carbon and oxygen, 1 to 5 ppm each of nickel and iron, and about 1 to 2 ppm of copper, with the remaining residual impurities being less than 1 ppm. Further indication of the purity of this material is provided by the critical resolved shear stress for basal slip, which was approximately 300 psi. The starting material for the alloy single crystals was 1-in.-diam floating-zone-refined (six passes) rod of Pechiney SR grade beryllium. Two such rods were wrapped respectively with sufficient weight of wire of high-purity copper (99.999 pct) or nickel (99.999 pct) to yield a 5 wt pct alloy. A seventh floating-zone pass was then applied to each of the rods to accomplish the initial alloying and an eighth pass for homogenization. Analytical samples were taken from regions of the rod immediately adjacent to where the mechanical test specimens were cut; these indicated 4.37 wt pct Cu and 5.24 wt pct Ni. 2) Crystal Orientation. To avoid the occurrence of basal slip during c axis compression testing, it is necessary to load the crystals as nearly parallel to the c axis as possible. Preliminary c axis compression tests indicated that plastic flow and/or fracture occurred at stresses of the order of 300,000 psi; hence on the basis of a critical resolved shear stress for basal slip of 300 to 400 psi, the maximum crystal misorientation permitted is about 4 to 5' of arc. Since this accuracy cannot be obtained using the usual back-

Jan 1, 1969

By I. A. Ibrahim, A. K. Ismail, A. A. I. Afifi

"Synthetic rutile for paints industry is conventionally produced by a combined smelting and leaching process. Smelting process is an energy intensive step and represents economic burden in view of the increasing energy costs. Egyptian ilmenite ore containing 40% TiO2 and 38% total Fe was thus directly leached with sulphuric acid to produce eventually synthetic rutile. The process-involves digestion of the ore with sulphuric acid whereby both TiO2 and Fe are dissolved. The solution was then treated with iron powder to convert Fe3+ to Fe2+ to avoid co-precipitation of iron during the hydrolysis step. After solid/liquid separation and clarification, Fe was crystallized as ferrous sulphate hydrate by chilling. The titanium pregnant liquor was concentrated by evaporation and then hydrolyzed to precipitate hydrated titanium dioxide which was then calcined to produce synthetic rutile containing 98.6 TiO2 and 0.58% Fe2O 3. The optimum conditions of all the processing stages were determined.IntroductionSynthetic rutile for paints, welding rods and other industries could be produced from ilmenite ore by sulfuric acid leaching process. This process is a batch process and nearly 80% of all TiO2 produced in the world is made by this process [1]. Figure l shows simple block flowsheet of the process. The process involves the following stages.Digestion: In this stage, ilmenite ore is treated with concentrated sulfuric acid where both iron and titanium are converted to their respective sulfates according to the following equations:"

Jan 1, 2000

By J Mauk, R Ruddock

The Taharoa ironsand deposit is 1600 hectares in size and is located 144 km south of Auckland on the west coast of the North Island, New Zealand. The sand predominantly contains heavy mineral-rich sediments, quartz, and feldspars. One hundred and two samples were collected from all areas of the mine, from the stockpile, black sand areas, tailings and Te Akeake Sands. Samples were collected with the primary focus on the next predicted mining path and secondly on gaining a good geographic spread throughout the mine. All samples were analysed for the dry, wet and particle density, the porosity and the magnetic, non-magnetic and clay minerals content.Stockpile samples had an average magnetic mineral content of 92 percent. Dry, wet and particle density values for stockpile material exceed all values from untreated ironsand samples with average values of 2.8g/cm3, 2.9g/cm3 and 4.4g/cm3 respectively. Black sand samples from the mine came from various areas such as the central area, the forest, the southern area and the benches. All black sand samples are remarkably consistent, with dry and wet density values between 2.3 û 2.5g/cm3 and particle densities that range from 3.7 to 3.9g/cm3. The porosity values are comparable to all other areas of the mine (average 37 per cent). The tailings and Te Akeake Sands have different physical properties due to their different mineral contents. The dry, wet and particle densities are all lower than that of all other areas with averages of 2.1g/cm3, 2.2g/cm3 and 3.3g/cm3 respectively. These lower values reflect the lower magnetic mineral content (average 18 per cent) and the associated increase in non-magnetic mineral content (average 78 per cent).

Jan 1, 2002

By J. A. Oraniscak

Longwall mining operations continue to increase extraction rates and the potential to generate higher amounts of airborne respirable dust (ARD). The National Institute for Occupational Safety and Health (NIOSH) has gathered data from underground longwall surveys and laboratory crushing tests to identify the important factors influencing the amount of ARD levels generated. Coal-seam characteristics, operational parameters (i.e.,cut sequence, mining height, etc.), extraction rates and dust-control parameters (i.e., airflow, water flow, etc.) were examined to determine their influence on ARD levels measured at longwalls during actual mining operations. ARD generated by a laboratory crusher while crushing bituminous coal samples collected from underground longwalls was also examined to determine if there were cor relations with coal-seam properties. Results from the longwall and laboratory data collected indicated that low-ash, high-volatile coals (low moist fuel ratio or MFR coals) generate higher amounts of ARD. It was found that face ventilation and water application to the coal product are essential elements to controlling longwall ARD levels.

Jan 1, 2000

By B. W. Frazier

IN the numbers of the Engineering and Mining Journal of June 27th and July 11th, 1874, there appeared some very complete statistics of the working of some anthracite blast-furnaces belonging to a large establishment in this country, followed by calculations of the heat requirements of the furnaces, similar to those which have been made for European furnaces by Mr. I. Lowthian Bell, Prof. Gruner, Prof. Akerman and others. To the author of these valuable papers, Mr. John A. Church, of New York, the thanks of American iron metallurgists are due for this attempt (the first, to my knowledge at least), to attack systematically and scientifically the problem of economy of fuel in our anthracite blast-furnaces, by applying the formulas obtained from European practice and from the experiments of European metallurgists, to the determination of the amounts and the modes of distribution of the heat production and absorption in our furnaces. In order to facilitate a comparison of the results thus obtained for the anthracite furnaces in question with those obtained for the best types of European coke furnaces, I have reduced to the same standard of heat units (the product of the ton and the degree Centigrade), and placed side by side, the calculations of the heat requirements of one of the furnaces of the Clarence works, as given by Mr. Bell, of a furnace at Pouzin, on the banks of the Rhone, given by Prof. Gruner, and of the anthracite furnaces for 1869-73, as given -by Mr. Church. Table I shows the different conditions of working of the three furnaces, of which Table II gives the respective heat requirements. I have made this latter table in such a manner as to give for each furnace its heat requirements, both according to the views of Mr. Bell and those of Prof. Gruner. Mr. Bell assumes that all the carbonic acid produced by the decomposition of the carbonate of lime of the charge, is reduced to carbonic oxide by contact with solid carbon ; also, that the carbon of the pig iron is a portion of that deposited in the ore in the upper part of the furnace, by the dissociation of carbonic oxide. Prof. Gruner, on the other hand, assumes that all the carbonic acid of the carbonates of the charge escapes unreduced, and that the carbon deposited in the upper part of the furnace, by the dissocia¬tion of carbonic oxide, is again converted into carbonic oxide or

Jan 1, 1875

By Michael Stelter

Lead alloys are a typical material for anodes in technical electrowinning processes. The life time of those anodes differs from 2 years to more than 5 years in tankhouse practice depending on the type of alloy, the mechanical damaging and last but not least the chemical corrosion in the electrolyte. Different types of alloys, such as PbSb, PbCa, PbAg, PbSnCa or PbAgCa have been developed and used in the past, their rate of corrosion being more or less a philosophic discussion. Comparable investigations on the corrosion behavior in dependence of the concentrations of alloying metals, the grain structure and the mechanical and thermal treatment during production have not been conducted before. In the last years the Institute for Nonferrous Metallurgy and Purest Materials of the TU Bergakademie Freiberg investigated the corrosion behavior of the typical lead alloys for metal electrowinning under comparable conditions. Correlations between different alloys, mechanical treatments like rolling or thermal treatment like annealing to the rate of corrosion were made, leading to the conclusion that grain structure and grain size are of eminent importance for the rate of corrosion in technical electrolytes. From these results new and optimized lead alloy anodes can be produced with an increase in life time of the factor 3 ? 5. The optimized lifetime in terms of corrosion stability can not protect from mechanical damaging in tankhouse practice, but with a careful practice the increase in lifetime can save a lot of money and problems.

Jan 1, 2006

By O. C. Wheeler

The reports for the years 1942 and 1943 were written in 1943 and 1944, respectively, and were summarized at the meetings of the Institute in February of those two years, but the manuscripts were held until released for publication in 1945. 1942 Lack of transportation facilities during 1942 resulted in a sharp curtailment of producing operations in Colombia at the middle of the year, and production of crude for export was thus greatly affected. Total production in 1942 amounted to only 10,617,008 bbl. Of this amount the Tropical Oil Co. produced 9,453,995 bbl., which includes 126,607 bbl. of petroleum condensate, and the Colombian Petroleum Co. produced 1,163,013 bbl, from its Barco property. The following summary, showing a comparison of the number of barrels of crude produced during the years 1941 . and 1942, reflects the extent to which producing operations were reduced during 1942: These above figures do not include any production recovered from Condor's (Shell's) Casabe wells during testing operations. An active exploratory drilling Program was carried out by several of the major companies, and the year 1942 established a record for wildcatting activity in Colombia. Excluding development operations on the Shell's Casabe wells on the Yondo concession, 13 wildcats were completed. Eight of these were in the Magdalena Valley, and all were abandoned; five were on the Barco Concession, and of these four were completed as producers. Particularly notable among the petroleum developments during 1942 was the continued success of drilling operations on the Shell's Yondo Concession of Compania Colombiana de Petroleos El Condor opposite Barranca-Bermeja. In 1941 the discovery well, Casabe No. I, was drilled and completed with an estimated initial production of approximately 480 bbl. of 19 1/2 A.P.I. gr. oil per day. In 1942, seven stepouts from the discovery well were successfully completed with initial production ratings ranging from 150 to 1230 bbl. per day. An extensive wildcat program on the Barco Concession of the Colombian Petroleum Co. yielded favorable results with the completion of four producers in the Socuavo and Tres Bocas areas; a fifth test, Sardinata Norte No. I, was abandoned at 500;" feet. Wildcatting operations in the lower Magdalena Valley area were initiated in 1942 With the drilling of a deep test, El Dote No. I, by Compania de petroleo Shell de Colombia. The well was a failure and was abandoned at a depth of 10,046 (?) feet. Wildcats on which drilling operations were being conducted at the end of the

Jan 1, 1945

By W. M. Courtis

In compliance with the suggestion in Mr. Emmons's paper on the Genesis of Certain Ore-Deposits (Trans. xv., 125), that facts should be accumulated for study and future compilation, I wish to put on record, before the data are lost, a section of the Animikie series of rocks and the vein-phenomena exposed by the explorations at the Duncan mine, Port Arthur, Thunder Bay, Lake Superior. This may serve as a supplement to my paper on this district (Trans. v., 473). I trust this information will not be without interest at. the present time. The recent discoveries of very rich silver-ores at Rabbit and Silver Mountains, on the same belt, have brought the district once more before the public; and a study of the facts here set forth may assist some fellow-engineer in the determination of the value of property offered for sale in this neighborhood. The vein at the Duncan mine was tested to a depth of over eight hundred feet by shaft and drifts, and to nearly a thousand feet with

Jan 1, 1887

By T. V. Jayaraman

Fe-Ga alloys, with compositions around 27.5 at % Ga, can be heat treated to obtain ? (ordered bcc), DO19 (ordered hexagonal), and L12 (ordered fcc) ordered structures. Our earlier work using polycrystalline samples had shown that magnetostriction is strongly influenced by ordering in this alloy. Here more detailed study of the correlation between ordering and magnetostriction using single crystals of Fe 27.5 at. % Ga alloys is reported. Single crystals of Fe 27.5 at % Ga were grown using Bridgman single crystal growth technique. X-ray and electron diffraction were used to study the ordering in this alloy. Magnetostriction measurements were carried out at no load conditions. Elastic constants were measured using Resonant Ultrasound Spectroscopy.

Jan 1, 2006

By T. P. Lim, A. J. Vivyurka, N. K. Davé

Chemical and radioisotope stratification studies were carried out in an abandoned uranium tailings pile at Elliot Lake, Ontario, Canada. The tailings, with an average thickness of 7 m over an area of 16 hectares, were deposited approximately 20 years ago on a layer of black peaty material which overlays deposits of permeable glaciofluvial sands. Geohydro-chemical investigations of this tailings area have been ongoing since 1978. The radioisotope profile study was conducted on three adjacent sites. Solid core samples, collected with a split spoon sampler at various depths, were analyzed for various chemical and radioisotope constituents. The results indicated by 2 distinct zones of stratification; an upper zone, where the constituents were uniformly distributed, and a lower zone, near the peat-tailing interface, which showed increasing concentration with depth. Within the peat layer, the concentration decreased rapidly with no significant levels of contaminant penetration below the peat layers observed, though the samples were limited. The pH profile showed that the top 5 m of the tailings have been oxidized, the average soil pH in this zone was 3.4 compared to 5 .8 underneath it. The groundwater hydrology of the system was observed to be complex showing downward and lateral flow in the dry season and lateral and upward flow during the spring and periods of heavy rainfall. The results are discussed in terms of a model based on the combined segregation and settling process of the tailings fine fraction, and the geohydro-chemical interaction of the solid and liquid phases.

Jan 1, 1981

By A. S. Baillie

THE subject of taxes appears to engage our attention altogether too lightly. At least it would seem that, notwithstanding the increase in the cost of government, and its concomitant increased taxation, little constructive work is being done with a view to solving the problem of increased taxes. I believe that the inattention of the public generally to the subject of increased taxation is due largely to the fact that shareholders are not advised of the amount of taxes paid by industry. Were companies to advise their shareholders the amount of monies paid or payable out of each year's earnings for taxes in one form or another, and were such amounts compared with the amounts available for distribution to shareholders, the effect would be a general demand for more equitable taxation methods. While I am not unmindful of the burdens and inequities in taxation confronting industry generally, I believe that the mining industry in this Province does not fully appreciate the burden represented by the costs of taxation. My reason for this belief is based partly upon1the fact that no serious effort has so far been made to devise some means By which the mining industry in British Columbia might protect itself against the ever-increasing burden of taxation. In addition to this fact the widely diverse nature of mining in British Columbia makes the accumulation of comparable data a problem which, so far as I can ascertain, has never been attempted in a manner which would be authoritative. It should not be necessary to cite the effects of this burden upon any one mining company or upon any branch of the mining industry, such as coal mining, placer mining, base-metal mining, or the precious-metal mines. In my opinion, the operation of our mines by large corporations is conducive to the payment of taxes with much less protest than would be the case were the taxes being taken out of the earnings of just a few partners or a very few shareholders. We appear to be too unmindful of the taxes, paid by mining companies, of a kind that are not so prominently brought to our attention as are the taxes which are calculated upon the taxable income of the company at the end of its fiscal year. British Columbian mining companies pay many other taxes besides those which are calculated upon taxable income. Among these are school taxes, property taxes, gasoline taxes, mineral taxes, coal royalty, to say nothing of fuel-oil tax, sales tax, and excise taxes.

Jan 1, 1935

By H. W. III Straley

SOME ten years ago the writer1 made a survey of college catalogues to determine what sort of training geologists were receiving in basic sciences. In the light of this compilation and subsequent experience, he made the statement at the 1939 meeting of the Institute that apparently only in schools offering a curriculum in geological engineering or in the geological option of mining engineering curricula could geologists hope to receive the training advocated by Hubbert.2 In the former study, schools of engineering were excluded because the preparation was known beforehand. The present study aims to consider engineering schools in an effort to determine not only what basic studies are required but what use is made of them in geological courses of the third and fourth years. The present study combined the method of consulting catalogues with a questionnaire. Letters were mailed to officials responsible for supervision of mineral industries curricula requesting information on: (I) the extent to which the engineering preparation of the first two years is actually used in the work of the last two; (2) the extent to which the engineering of the last two years is used in courses in geological engineering or geology; and (3) whether geologists or geological engineers teach descriptively or analytically. Both methods of securing information are open to criticism; the former on the basis of the incomplete information supplied by university catalogues concerning prerequisites and curricula and the latter on the ground that the questions may be phrased poorly or given insufficient consideration by replying officials. Fifty-six questionnaires were mailed and 39 (70 per cent) replies were received. Of the replies, 6 (11.3 per cent) were from colleges of liberal arts unconnected with engineering schools, 11 (20 per cent) were from schools having curricula in geological engineering and 22 (39 per cent) were from departments of mining engineering. The number of replies and their distribution were satisfactory in spite of the receipt of only six from liberal arts colleges. It should be understood that the universities of which the catalogues were examined were not necessarily the same in both compilations. No departments that are admittedly or known to be acting solely as service departments were recorded in the second study. Table I shows that all professional schools require mathematics through calculus, whereas only 7 per cent of the liberal arts curricula expect the same level of preparation; only 73 per cent require trigonometry and 40 per cent college algebra. The proportion of geological engineering curricula requiring differential equations is higher than mining. All of the engineering schools require general and analytical chemistry, whereas only 77 per cent of liberal arts curricula require general and 18 per cent analytical. Twenty-two per cent of the mining and 40 per cent of the geological engineering

Jan 1, 1941

By C. E. McCarthy, A. George Stern, Stanley J. Green

The national interest prompts consideration of any new source of mineral wealth even though the immediate need may be of minor importance. A critical shortage of potash in the United States during the World War of 1914-18 stimulated broad development, which established an American potash industry on a basis so sound that this commodity is not even listed as urgent in the present war program. However, this should not minimize its importance, and the need remains, not only for adequate supplies of the usual potash salts (the chloride and the sulphate) but of all potash chemicals. Potassium carbonate (known commercially as pearl ash) has been a high-cost commodity; in consequence, the quantity used by industry has been relatively small. By a combination of fortuitous geographical circumstances, there is opportunity to produce a low-cost potassium carbonate with the possibility that this might become an important source of chemical potash. The following paper considers a recently developed process1 taking advantage of these conditions. Wyomingite Deposit For many years it has been known that a deposit of potash-bearing igneous rock known as wyomingite occurs in the south- western part of Wyoming at Zirkel Mesa, which is 44 miles northeast of the town of Green River (Fig. I). The western escarpment of this deposit is about 3 miles east of the Superior Branch of the Union Pacific Railroad. This mesa alone covers approximately 3300 acres, and the average thickness of the wyomingite flow is 75 ft. It has been estimated2 that about one billion tons of wyomingite occur in this one deposit, all of which can be mined by open quarrying without overburden. In addition, there are several smaller deposits within 10 miles, so that the aggregate wyomingite in the area approaches two billion tons. This is one of the world's largest sources of leucite, and the only known deposit of its kind in the United States. The wyomingite (which is in the form of a lava flow and numerous small craters), with the mineral leucite as its most important component, has a mineral composition3 of: leucite, 50 per cent; phlogopite, 15; diopside, 12; kataphorite, 15; glass and apatite, 8. The phlogopite is present in visible phenocrysts 1/8 to 1/2 in. in diameter, but the leucite usually is almost microcrystalline and not capable of liberation by grinding. Chemically, the Wyomingite has a potash Content averaging slightly higher than eleven per cent K2O. The potash is mainly in the leucite, but part of it is in the phlogopite and a small amount is in the feldspars that are occasionally seen as undigested occlusions of the granite country rock, through which the leucite-bearing lava intruded, and as sanidine feldspar both in the crystalline and glassy phases of the rock. A typical analysis of the wyomingite from Zirkel Mesa is as follows:

Jan 1, 1948

By A. George Stern, Stanley J. Green, C. E. McCarthy

The national interest prompts consideration of any new source of mineral wealth even though the immediate need may be of minor importance. A critical shortage of potash in the United States during the World War of 1914-18 stimulated broad development, which established an American potash industry on a basis so sound that this commodity is not even listed as urgent in the present war program. However, this should not minimize its importance, and the need remains, not only for adequate supplies of the usual potash salts (the chloride and the sulphate) but of all potash chemicals. Potassium carbonate (known commercially as pearl ash) has been a high-cost commodity; in consequence, the quantity used by industry has been relatively small. By a combination of fortuitous geographical circumstances, there is opportunity to produce a low-cost potassium carbonate with the possibility that this might become an important source of chemical potash. The following paper considers a recently developed process1 taking advantage of these conditions. Wyomingite Deposit For many years it has been known that a deposit of potash-bearing igneous rock known as wyomingite occurs in the south- western part of Wyoming at Zirkel Mesa, which is 44 miles northeast of the town of Green River (Fig. I). The western escarpment of this deposit is about 3 miles east of the Superior Branch of the Union Pacific Railroad. This mesa alone covers approximately 3300 acres, and the average thickness of the wyomingite flow is 75 ft. It has been estimated2 that about one billion tons of wyomingite occur in this one deposit, all of which can be mined by open quarrying without overburden. In addition, there are several smaller deposits within 10 miles, so that the aggregate wyomingite in the area approaches two billion tons. This is one of the world's largest sources of leucite, and the only known deposit of its kind in the United States. The wyomingite (which is in the form of a lava flow and numerous small craters), with the mineral leucite as its most important component, has a mineral composition3 of: leucite, 50 per cent; phlogopite, 15; diopside, 12; kataphorite, 15; glass and apatite, 8. The phlogopite is present in visible phenocrysts 1/8 to 1/2 in. in diameter, but the leucite usually is almost microcrystalline and not capable of liberation by grinding. Chemically, the Wyomingite has a potash Content averaging slightly higher than eleven per cent K2O. The potash is mainly in the leucite, but part of it is in the phlogopite and a small amount is in the feldspars that are occasionally seen as undigested occlusions of the granite country rock, through which the leucite-bearing lava intruded, and as sanidine feldspar both in the crystalline and glassy phases of the rock. A typical analysis of the wyomingite from Zirkel Mesa is as follows:

Jan 1, 1948

You have been involved with research activities in froth flotation and size reduction. What do you think the future holds for these technologies? Froth flotation and size reduction are essential processing steps common to almost all current mineral-processing operations. The most significant recent advances in both of these technologies have been in the increased size of equipment and in some alternative machine designs, such as the column cells and the high-pressure roller mills. I believe that the next major area of improvement in size reduction will be at least partially driven by an improved awareness of controlling the conditions that influence the efficiency of particle breakage. This would include factors such as water-chemistry influences, the influence of fines from classification, chemical grinding additives and better rheology control. In froth flotation, significant changes are currently taking place in the plant use of frothers. Also, there will be an increasing number of new collector chemistries for certain types of flotation, such as with oxide minerals, oxidized mineral recovery in sulfide systems and improved selectivity with certain sulfide mineral systems. You have been active in implementing new engineering technology in industry. What do you think will be the technological future of mineral processing? Being involved in the invention and industrial implementation of new technology has been one of the most positive experiences of my career. I also have had the unique experience of working in the chemical and petroleum industries as well as in mineral processing. In my opinion, the future of mineral processing will be an exciting, wonderful work area in which to be involved. The development of new technology in minerals processing offers a special challenge compared, for example, to that of chemistry or chemical engineering. In successful industrial innovation involving minerals, having a sound theory is not enough. Instead, what is required is a valid blend of a good idea, meaningful laboratory characterization and, most key, the experience necessary to adjust the plant in such a way to incorporate the idea. The complexity of engineering systems that contain solids, liquids and gases, as well as highly variable natural mineral properties, means that nothing new comes easily. In some cases, it has been difficult to explain how processes practiced in industry for decades actually work, in terms of physical and chemical fundamentals. On the other hand, when a new idea is successfully implemented, the economic and professional returns can be very rewarding. I know of no more challenging field of engineering than minerals processing. What is your view of the future of the minerals industry in general? The need for minerals, metals and coal will be with us for decades to come. In many aspects, the minerals industry became global much earlier than some of the other industrial sectors that are currently experiencing globalization. I believe that fair international competition is beneficial to the users of minerals, metals and coals and to those involved in producing such materials. The key word here for a global economy to work is "fair." So the various playing fields involved such as finance, profit return, regulation and labor should be relatively level for all the players involved. Competition is an amazing driving force for innovation. This has been true in the minerals industry during the last few decades. The mining industry is producing more at less cost per unit than ever before in our history. I anticipate that this trend will continue as long as the global political situation remains relatively stable.

Jan 1, 1997