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By S. G. Malghan, A. L. Mular

Size distribution and surface area measurements play an important role in the characterization of crushed and ground granular materials properties. This paper outlines sampling methods of granular materials followed by a detailed treatment of principles of size and surface area measurements of a wide range of sizes from crushers feed to ball-mill ground sub-micron particles. Currently available equipment for size distribution and surface area analysis is described and critically examined with respect to advantages, range of application and precautions in their use. Equipment for on- stream size analysis is also evaluated. Finally, a list of criteria are given for selection of particles size and surface area measuring devices for a particular application.

Jan 1, 1982

By H. C. Theuerer

EVEN the highest purity silicon available for semiconductor use contains significant amounts of donors and acceptors, usually aluminum, phosphorus, and boron. Aluminum and phosphorus can be removed from silicon by adaptations of the zone refining techniques' widely used with germanium. Removal of boron, however, has been a difficult problem. The approximate distribution coefficient for boron in silicon is 0.8, so that while zone refining is possible, it is not very effective in any reasonable number of passes. It has been found, however, that boron can be effectively removed from silicon by treatment of a liquid zone with hydrogen and water vapor. Starting with 40 ohm-cm, n-type silicon, use of zone refining coupled with hydrogen

Jan 1, 1957

By G. K. Manning, A. R. Elsea, C. R. Simcoe

It was found that a critical boron content exists which yields the maximum boron-hardenability effect in hypoeutectoid steels, as was predicted from the mechanism proposed in a previous paper. The hardenability of boron steels with boron contents between 0.00005 and 0.0017 pct was dependent on the boron content, austenitizing temperature, quenching temperature, and austenite grain size. The effects of these variables on hardenability can be explained by the mechanism previously proposed for the boron effect. A MECHANISM has been proposed to explain the effect of boron on the hardenability of steel.' The proposed mechanism assumes that transformation particles nucleate at concentrations of lattice imperfections where the energy level is high, and

Jan 1, 1957

By George R. Eadie

As 1972 slipped through our grasp, coal industry adjustments to health and safety laws, environmental restrictions, rising costs and higher production targets left many mine operators groping for a new grip. The industry adopted a defensive attitude in 1972 as it faced the onslaught of inspectors. With rigid demands for improving safety and instituting environmental programs, the industry worked hard and watched production figures stagger upward while operating costs zoomed. In an effort to determine how to meet the goals that were established for fewer ignitions, fewer accidents, less respirable dust, lower noise levels, fewer law violations, better reclamation, less pollution, lower costs and higher production, industry representatives attended numerous technical meetings.

Jan 2, 1973

By N. B. Davis

I . INTRODUCTION WHILE working with a number of very sticky cracking clays from western Canada the writer became interested in a study of the cause of the excessive plasticity, This led to a review of the rather bulky literature on the subject and the finding that plasticity still remains to be satisfactorily explained. Most of these investigations seem to have been carried on solely from the chemical viewpoint, the geologic aspects having been ignored. In nearly every case, the investigators have worked with the higher-grade clays of low to medium plasticity, such as the kaolins, the excessively plastic ones having generally been neglected, partly because of their limited application, yet the author believes that these are

Jan 2, 1915

With one shaft almost completed and another shaft being sunk, the scope of the Saskatchewan potash area is beginning to come into focus. Some 18 companies have reportedly leased more than four million acres to explore potash possibilities. PCA, farthest along, is located about 15 miles southeast of Saskatoon, and IMC has its project underway near Esterhazy in the eastern edge of the province. U. S. Borax and Chemical Corp. is also reported exploring the area. The field lies in the middle Devonian geological formation which was probed by oil companies along the 200-mile area across the province. Extensive examination began after World War II and deep drilling was started a few years later.

Jan 5, 1958

By Louis W. Huber

The Liberty tunnels extend through a very steep hill in Pittsburgh (locally called Mount Washington) for a distance of slightly over a mile. The two tunnels parallel each other and are 59 ft. apart, center to center (Fig. 1) They are for vehicular traffic, and in each there is ample space for two lines of vehicles. Northbound traffic, toward downtown Pitts- burgh, is confined to the east tunnel and southbound traffic to the west tunnel. The South Hills residential district is separated from downtown Pittsburgh by Mount Washington, a steeply rising escarpment south of and adjacent to the Monongahela River. The Mount Washington district immediately south of the Monongahela River has been reached, for many

Jan 1, 1927

By Richard F. Moe

INTEREST in developing White Pine, considered since 1942, was renewed by the Korean conflict and its shortage of domestic sources of copper. In view of this Morris F. La Croix, president of Copper Range Co. of Boston, and of its wholly owned subsidiary, White Pine Copper Co., worked an agreement out with Government officials. Essentially the contract called for a construction loan of $57,185,000 at 5 pct interest, with $13 million additional supplied by Copper Range. Defense Materials Procurement Agency contracted for 243,750 tons of copper from the first 275,000 tons produced, at a base price of 25.5& per lb, the contract to expire Dec. 31, 1961.

Jan 4, 1954

By W. C. Phalen

The production of salt in the United States divides itself at the outset into two distinct classes…

Jan 1, 1915

By G. T. Smith, R. C. Moyer

Cadmium, along with other impurities such as lead, gallium, germanium and indium, is characteristically found associated with zinc ores, the average ratio of zinc to cadmium being about 200 to 1. The increasing demand for cadmium has led most zinc producers to develop a means for its recovery. The separation and recovery of this byproduct constitutes an important development in the metallurgical field. Zinc sulphide concentrates, principally from domestic sources, are processed at Donora Zinc Works and they assay approximately 60 pct zinc, 30 pct sulphur, 0.37 pct cadmium, and 0.8 pct lead. Concentrates are roasted to convert zinc sulphide to zinc oxide. Sulphur evolved as sulphur dioxide is used to manufacture 60º Baume sulphuric acid by the lead chamber process. The roasted calcine is sintered to remove the remaining sulphur and volatile impurities. Fume from the sintering machines is collected by Cottrell pre-cipitators and sent to the Cadmium Department for cadmium removal. The sintered product, a crude zinc oxide, is then mixed with coke and salt and charged into horizontal retorts for reduction to zinc metal. Furnace residues are then passed through a Waelz kiln which recovers practically all of the remaining zinc as an oxide. This oxide is collected by Cottrell precipi-tators and returned to the sintering plant for further processing. The following discussion covers operating procedure, and equipment used in the Cadmium Department for the manufacture of cadmium and its byproducts. (See attached Flow Sheet, Fig 1.) The production of cadmium may be divided into several steps, each of which will be discussed separately. These steps are as follows: 1. Raw Materials—Sinter Plant Cottrell dust, acid sludge and sulphuric acid. 2. Sulphating—Separation of lead. 3. Purification—Separation of impurities other than zinc. 4. Precipitation—Separation of cadmium from zinc. 5. Briquetting—Preparation of sponge for charging to furnace. 6. Smelting—Distillation of cadmium metal. 7. Remelting—Casting into commercial shapes. 8. Recovery of Rare Metals—Separation of germanium, gallium and indium. Raw Materials The raw material, or Sinter Plant Cottrell dust, from which the cadmium is manufactured, is recovered by Cottrell units collecting fume from two Dwight-Lloyd sintering machines. An average analysis of this dust is 36 pct zinc, 1.5 pct sulphur, 17 pct lead, and 5.5 pct cadmium, with associated elements in smaller amounts. A lesser tonnage of material removed from the sintering machine fans, and called fan cleanings, is also treated for cadmium. It is lower grade material and analyzes approximately 41 pct zinc, 1.2 pct sulphur, 14 pct lead, and 4 pct cadmium. Cottrell dust and fan cleanings are delivered to the Cadmium Department by monorail crane in bell-bottomed dump buckets. Acid sludge and 60" Bauine sulphuric acid are received from the Acid Department. The acid sludge is a mixture of lead sulphate, sulphuric acid and zinc sulphate, and is recovered mainly from the floors of the acid chambers or storage tanks. The acid sludge is delivered to the Cadmium Plant in a lead lined bucket. The 60º Baume sulphuric acid is delivered through a lead pipe to a lead lined storage tank. Sulphating Four well ventilated lead lined sulphating tanks, 8 1/2 ft in diam by 12 ft deep, are used. These are steel tanks lined with 20 lb lead on the bottom and 16 lb lead on the ides, equipped with Hastelloy side agitators. 15 hp motors drive 24 in. propellers at 450 rpm. There are water connections to stuffing boxes to keep sludge from entering the bearings. The tanks are covered with removable wooden lids and are connected to marine plywood

Jan 1, 1950

By A. L. Lee, M. H. Gonzalez, R. F. Bukacek

Experimental viscosity data for methane are presented for temperatures from 100 to 340F and pressures from 200 to 8,000 psia. A summary is given of the available data for methane, and a comparison is presented for that data common to the experimental range reported in this paper. Correlation of the data is presented, and predicted values are given for temperatures up to 460F and pressures up to 10,000 psia. INTRODUCTION The increasing ranges of temperature and pressure at which fluids are produced, transferred and processed in the petroleum and chemical industries stress the need for accurate information on physical properties, both for engineering calculation and for improving the methods used to estimate physical properties. A survey of the literature reveals that disagreements between published data on the viscosity of methane are common and that most investigations have been conducted over restricted temperature and pressure ranges. This situation made it desirable to undertake an investigation with a scope common to most of the literature data available. APPARATUS AND MATERIALS The apparatus used in this investigation was the capillary tube viscometer'described by Dolan,10 with modifications introduced in the general design and operation of the instrument. It has an effective pressure range from 14.7 to 10,000 psia and a temperature range from room temperature to 400F. The design of the viscometer is based on the establishment of a manometric head between two vessels containing the test fluid and a volume of mercury. The reservoirs are connected by a capillary tube through which the test fluid flows and a tube through which mercury flows. A pressure gradient is established by elevating one of the vessels above the other, the resulting flow of mercury displacing the test fluid through the capillary. The schematic diagram of the system (Fig. 1) shows the arrangement of the equipment auxiliary to the viscometer. The density cells assembly (E, Fig. 1) was not used in this investigation since reliable data are available on methane. 23 The assembly consists of a bank of eight 316 stainless steel pycnometers. Experimental values of the density of a natural gas19 and i-butane14 samples have been obtained for pressures up to 8,000 psia and temperatures up to 340F. The methane used was obtained from the Southern California Gas Co. Mass spectrometric analysis showed a composition of 99.8 percent CH4, 0.1 percent C2H6 and 0.1 percent C3H8. EXPERIMENTAL VALUES AND ACCURACY OF DATA Determinations were carried out at several temperatures. With the temperature in the cabinet controlled at a given level, a series of runs were made at various pressures. To establish pressures lower than that obtained from the methane cylinder, it was necessary to vent the gas slowly to the atmosphere until the desired pressure was reached. The consistency of the test run time for a particular condition can be used as a criterion for the accuracy of the data. A further check is obtained by comparison of the viscosity values obtained using different driving heads. To establish the instrument's reproducibility, a large number of runs were made at three completely different conditions of temperature and pressure. Data obtained were analyzed statistically to determine the most reliable value for the variance of the experimental data. It was found that the variances of the data taken at different experimental conditions were not statistically equivalent, but they were small. The largest variance was found at 8,000 psia and 100F, where the standard deviation was 0.385 micropoise. Therefore, 99.9 percent of the data can be expected to fall within 1.15 micropoises of the population mean, except for the intervention of systematic errors in operation.

By Michael A. Gustafson, Fernando B. Sotelino

INTRODUCTION The degree of success of any new project will ultimately depend on two factors: (i) the underlying economic strength of the project; and (ii) how successfully the parties involved can coordinate their interests to establish a financing structure which indeed capitalizes on the project's inherent economic attractiveness. A project's underlying economic attractiveness is determined by its inherent comparative advantage in accessing the factors of production, combining these factors of production to create a product and marketing this product. Initially, the decision regarding whether to go ahead with the project - the go/no go decision - has to be based on its total ability to generate cash flow returns to project participants to compensate them for resources committed and risks undertaken. The extent to which project sponsors will actually benefit from the project's inherent economic attractiveness is fundamentally dependent on their ability to establish the long-term financing arrangement which best meets the objectives and constraints of each project participant and sources the most advantageous funding. The decision regarding the financing of the project, the "how-to-go" decision, is then an attempt to find the capital structure, consistent with market terms and conditions and sound financial management policies which will best: (i) match planned outflows with cash inflows; and (ii) maximize the return for project sponsors. This article makes use of a case study to illustrate the analytical process involved in dealing with the go/no go and the how-to-go decisions on a project. Though most commonly applied to project finance situations, the guideline described herein is equally applicable to other financing situations, such as leveraged buyouts and debt restructurings, where timely repayment of debt depends largely on the future cash flows to be generated by the entity being financed. Our case study is of a hypothetical expansion of an existing copper mining operation in Latin America. The company, Latin American Copper (LAC) currently has 147 million metric tons of proven sulphide ore reserves. The existing mine complex consists of an open pit mining operation and a conventional flotation plant with a processing capacity of 20,000 metric tons of ore per day. LAC sells copper concentrates primarily to Japanese custom smelters. The mine has been in operation for 5 years, fine copper production last year was 108,240 metric tons, and LAC currently has assets of $175 million and net worth of $135 million. Due to a recent increase in proven reserves, LAC is considering doubling the treatment capacity of its flotation plant to 40,000 tons per day, and expanding its mining capacity to meet the additional concentrator capacity (the "Project"). Project capital expenditures and incremental working capital have been estimated at approximately $210 million. The construction period is expected to cover one year, including a three-month start-up phase. The Project is to be funded by a combination of equity, senior debt, including both supplier credits and/or commercial bank term loans, and subordinated debt. LAC's shareholders are willing to commit a maximum of $50 million to the Project. A government development agency, which is also a minority shareholder, might be willing to lend approximately $40 million to LAC in the form of subordinated notes, depending on the ability of the Project to generate foreign exchange reserves. Commercial banks cannot be expected to offer loans with maturities longer than ten years. The Project is expected to increase the value of the firm because: (i) It will solidify the Company's position as a low-cost copper producer by taking advantage of economies of scale. As a result of the expansion, cash production costs are expected to decrease 7.0[c]/lb. and all-in costs 11.0[c]/lb; (ii) The potential earnings stream from the sulphide ore reserves will be realized

Jan 1, 1985

By Robert K. Waits

The resistivity, temperature coefficient of resistance, stability at 200°C, and structure of annealed Si-Cr films have been studied as a function of film cowposition. Colorimetric analyses of the films indicated 20 to 60 wt pct Cr over the resistivity range of 1 to 10-4 ohm-cm. Resistors were delinealed by photolithographic etching, contacted with aluminum, and mounted in TO-5 packages. The average temperature coefjicient of resistance, a, between 50" and 150°C is related empirically to the film resistivity, p, by a (ppm per °C) = -2900 - 1140 log p (ohm-cm). Near a resistivity of 3.5 x 10-3 ohm-ctn (30 at. pct Cr) , the temperature coefficient changes from positive to negative. The temperature at which the resistance maximum occurs increases with decreasing resistivity. The conduction mechanism between —196" and +360 C appears to have two components—an activated conduction dominating a1 lord chromium content and higher temperatures and a linearly temperature-dependent conduction dominating at high chromium content and low temperatures. Electron diffraction patterns showed the films to consist of Cr3Si and amorphous silicon. Resistors thicker than 150A typically drifted less than 0.6 pct after 5000 hr stor-age at 200°C. Resistors as thin as approxinzately 80°A that had a silicon overlayer showed similar stability. Such films had sheet resistivities of 10 kilohm per square and temperature coefficients near —600 ppm per "c. THIN-FILM resistors having sheet resistivities greater than 1 kilohm per square are required in many integrated circuit applications. A process for producing such resistors should be compatible with present circuit fabrication and packaging, including exposure to maximum temperatures of 300° to 570°C during contact alloying, die-attach, and ceramic package sealing. In addition, the resistors should have a low temperature coefficient of resistance and change less than 1 pct during storage at temperatures up to 200°C for 1000 hr or longer. Reports of work beginning in 1949 at Battelle Memorial Institute indicated that, among other materials, the silicides of chromium, Cr-Co, molybdenum, tantalum, and tungsten were worthy of further study as high-stability resistive films. Later work on Si-Cr films vacuum-deposited by flash evaporation from a tungsten boat1 led to the conclusion that Si-Cr had "excellent properties for development into high- stability ... resistive elements". More recently, Si-Cr films have been vacuum-deposited from a tantalum boat,2 from a BeO-coated tantalum boat,3 by flash evaporation from an electron-beam-heated tungsten Source,4 and by triode sputtering.5 This paper describes a study of Si-Cr films vacuum-deposited by electron-beam heating of Si-Cr alloys. Resistivity, temperature coefficient of resistance, stability at 200°C, and structure were studied as a function of film composition. EXPERIMENTAL PROCEDURE Film Deposition. Si-Cr sources were prepared by vacuum-melting pressed pellets of mixtures of powdered chromium (99.9 pct, -100 mesh, Lunex Co., Pleasant Valley, Iowa) and silicon (99.99 pct, -325 mesh, United Mineral and Chemical Co., New York). A permanent-magnet focused electron-beam gun having a water-cooled copper crucible was used to heat the pellet until complete melting occurred. The mixture was then kept molten for an additional 5 min. The alloy thus formed was used as the source material for Si-Cr film depositions using the same electron-beam gun. Pieces of semiconductor-grade silicon ingots were used as sources for silicon depositions. The films were deposited within a standard vacuum evaporator having an 18-in.-diam Pyrex bell jar, a liquid-nitrogen-cooled trap, and a Consolidated Vacuum Corp. 6-in. diffusion pump containing Dow Corning DC-705 pump fluid. The substrates were thermally oxidized* silicon wafers 2 to 2.5 cm in diam. The

Jan 1, 1969

By Maurice Lacasse, Pierre Legast

GENERAL OUTLINE For more than one hundred years, asbestos ore has been mined on Les Mines d’Amiante Bell's property at Thetford-Mines , Quebec, Canada. In 1878, the first tons of fibre were produced. The mine operated as an open pit until 1951 at a yearly production of about 330 000 mt of ore. Since 1951, the mine has operated underground using a block caving method. The mine produced 725 000 mt per year until 1978, when the rate of production was increased to 1 100 000 mt. The ore was produced entirely fm the first lift on 650 level until 1976. Then, a second lift on 950 level was put into operation to replace production facilities fm the first lift which is now exhausted. Annual fibre production is 72 500 mt with grades £ran 3 to 7. The total work force is 600 people. The host rock is a peridotite partially serpentinized. The asbestos fibre seams occur as stockwork deposit structurally controlled by talcose shear zone. The geological rock formations are heavily jointed and have a wide range of competence. Asbestos ore bodies have an heterogeneous nature which des difficult geological interpretation. THE GRIZZLY METHODS A full gravity method was used on the first level where the caved ore was drawn from finger raises onto grizzly bars spaced at 480 mn. Oversize blocks were reduced by secondary blasting. The ore fell by gravity in- to the ore pass below the grizzly chamber and was loaded into 5.7 mt haulage cars through a 1.2 m wide chute. Access to the undercut drift was fm the grizzly drift through finger raises. The undercut trough was drilled and blasted in a 1.2 m interval ring sequence into a slot situated at the end of each panel. The undercut stope was emptied through the finger raises. The undercut retreated from South to North at blast interval determined by the finger raises spacing. Fig. 1 represents the various steps involved with the grizzly method. Extraction drift dimensions were opened 1.8 m wide x 2.4 m high with anchor rock bolts and wire mesh where safety required them during excavation. Grizzly chambers were excavated at 2.4 x 2.4 x 4.6 m long on a 7.6 x 7.6 m grid spacing. Ore passes with dimensions of 1.5 x 1.5 m were developed from the haulage drift. On the extraction level, headings were drilled with jacklegs and cleaned by slusher into ore passes to the haulage chutes. The support used in the grizzly chambers was square concrete structure with 150 mn H beam reinforcement. The H beam was partially buried in the concrete and had one free face to allow for difference frequency of blast vibration between concrete and steel. The supported brow of the chamber had an opening 1.5 x 2.1 m. Ore passes were ring concreted 150 mn thick without any reinforcement. A 24 MPa compressive strength concrete was pumped and poured in pre-assembled forms. Based on the success of this method on the first level, the same layout was repeated on the second level. With the start of undercutting on the second lift, signs of deterioration were noticed in extraction drifts. At first, there was great concern as the operators were confident that standard repair techniques could cope with such a situation. An intensive program of support repair and a reduction of draw resulted. The following method of repairs were used: 1) Steel square sets of 150 mn H beam at 0.6 m intervals were installed. 2) This was reduced later to 0.3 m interval and eventually to a skin to skin steel structure. 3) Yieldable arches were introduced spaced at 0.9 m intervals, linked laterally with 150 m. H beam. 4) The yieldable arches spacing was then reduced to 0.46 m and eventually installed skin to skin. However, this support was not successful.

Jan 1, 1981

By Sidney G. Parker, Jack E. Pinnell

Cubic ZnSe crystals have been grown with HCl, HBr, and I, as chemical transport reagents. The growth of large, well faceted crystals is in the order HCl < HBr < I, with some produced by I, transport being 1 cm on an edge. ZnSe:I crystals showed no strain in cross polarized light and were homogeneous as determined by electron microprobe. These ZnSe:I crystals had a mobility at 25°C of about 300 sq cm per v-sec and contained 103 dislocations per sq cm; further, these crystals were pure cubic and showed both photolumi-nescence and photo conductance. Doping was attempted during crystal growth. ZnS and Zn (S,Se) crystals were also grown by I, transport. ZnSe crystals have been grown by most of the common techniques.1"6 Of these, chemical vapor transport as developed by Kaldis2 showed promise of producing high perfection, cubic ZnSe. With chemical transport, crystals can be grown far below the transition temperature for conversion of cubic to hexagonal structure. A small temperature gradient between source and growth region tends to allow near equilibrium growth, which should lead to good crystalline perfection. Because it is necessary to form volatile compounds, halogens or their compounds are extensively used as transport agents. HCl, HBr, and I, were compared as transport agents for growth of ZnSe crystals. Doping during crystal growth was attempted. ZnS and ZnSxSe1-x crystals were also grown by chemical transport. Electrical and optical properties of the crystals were examined. EXPERIMENTAL Fig. 1 is a sketch of the growth apparatus, which was a modification of Kaldis&apos;. A quartz rod which butted against the seed crystal served as a heat sink to promote growth only on the seed. Initially, crystal growth was attempted with ZnSe prepared by wet chemical methods. Because such a material contained oxides and selenates, HC1 reacted with these oxygen compounds to form H2O and no ZnSe transport occurred. Further, this ZnSe contained about 10 ppm Cu plus other impurities. High purity ZnSe, containing small amounts or no oxygen compounds, was prepared by reaction of H2Se with zinc vapors at 600° to 800°C in an atmosphere of palladium diffused H2.7 The ZnSe thus produced was usually copper free and contained only 1 ppm by weight of magnesium as determined by emission spectrographic analysis. G.E.&apos;s electronic grade ZnS, which had about 10 ppm metallic impurities, was used without further treatment. ZnS-ZnSe charges were prepared by mixing the two compounds together, then sintering 24 hr at 1100°C in an H2S ambient. The gases, HBr and HC1, were commercial grade, of 99.9 pct purity, while the I2 was 99.999 pct pure. The single crystal seeds with (111) faces for growth were mechanically held in place by sealing the outer quartz tube to the growth tube. After the charge was introduced, the system was outgassed at l0-5 torr while heated to 500°C. The desired quantities of transport reagents were then sealed off in the system, which had been cooled to 25°C. The growth tubes were inserted into a furnace with the desired temperature and thermal gradient. The system was allowed to cool for 1 to 12 hr from growth temperatures of about 850" to 25°C. The tubes remained stationary during crystal growth. HCl CHEMICAL TRANSPORT GROWTH Preliminary runs with HC1 had a temperature gradient of 100° to 200°C and no seed or heat sink. Only small crystals were obtained in these experiments as

Jan 1, 1970

By J. McLaren Forbes

Geologists and engineers in charge of diamond drilling often have no practical experience in the work, yet they must direct the drillers. In discussing some of the problems that arise, the author of this paper has tried to remember the things he did not know about diamond drilling when he first came in contact with that work. The basic reason for most diamond-drill holes is to obtain information. The more thorough the understanding of the geological and engineering problems involved, the greater will be the benefit derived from a drilling program. One of the best methods to ensure some thinking before drilling is to record on the permanent drill log, before the hole is collared, the reasons for the work and the information desired from that particular hole. When the hole is finished, the information obtained should be recorded. Driller The choice of a diamond driller is very important. If no experienced driller is available, a competent miner who has a sense of responsibility and better than average mechanical ingenuity can do the work. The best answer to most practical drilling problems, from the standpoint of the person in charge of the drilling program, is the service of experienced diamond drillers or drilling firms. Equipment Drilling equipment should be adequate to put down the longest hole desired and handle the toughest ground encountered. The type of bit used depends upon the ground drilled. Experimenting with various bits and careful study of footage records are the best means of determining which to use. Under ordinary circumstances it is best to drill using an EX barrel when possible. The length of rod and core barrels that can be used depends on the size of the drill station. Also, the length of core barrel depends on the way the ground cores. In good ground as much core will be recovered with a 10-ft. barrel as with a shorter one, and with considerable saving of drilling time. A water pump is needed if the hole is so deep that the water pressure is not strong enough to wash out the cuttings, or if it is drilled at an angle that prevents full force of the water pressure. Instructions For an Inexperienced Driller 1. Never run a bit too long a time. When a bit seems to be "good for one more run," discard it. 2. Always let the rods down on bottom easily. Protect the bit from sudden shock. 3. Start the drilling slowly and easily. Do not put pressure on the bit too fast. Diamonds can break. 4. Be sure the water pressure is strong enough to clear the bit, but do not use more water than is needed, as it may wash away the core. Usually, if no core is recovered the hole is a total loss. 5. When dry-blocking, do not run without water longer than necessary as the bit may burn. 6. Be sure the rods are well caulked and tightly screwed together. If not well caulked they will lose water, and if not

Jan 1, 1946

By William Romanoff

As you are no doubt aware, virgin copper, tin and zinc are strategic metals today and little, if any, is available in the manufacture of brass and bronze ingots. The source of alloys used in the manufacture of implements of war are, therefore, from ingots made almost exclusively from the refining of available scrap. It is amusing to hear of the unusual physical properties of various alloys because they mean nothing unless the foundryman is able to produce good sound castings. A discussion therefore means little unless something is said about the relative foundry practice of the various brass and bronze alloys. Among the factors that influence the quality of castings produced are: (I) improper gating, (2) wrong pouring temperatures, (3) poor sand conditions, (4) bad melting practice, (5) faulty pattern equipment, (6) metal. It seems to be the consensus of opinion that about 95 per cent of defective castings can be attributed to the first three causes. Metal while a source of trouble at times, is probably the least responsible for foundry difficulties. As you all know, a molder never makes a bad mold. His first thought is to blame the "damn" metal as this is the path of least resistance. Let&apos;s consider the "damn" metal and how it is made. In the manufacture of ingot, scrap is collected at various sources and is shipped to the refiner either directly or through the scrap dealer or broker. It is usually sorted, although not always. In working toward a specified chemical analysis, weighed amounts of various grades of scrap metals are melted, and then oxidized by blowing air through the molten bath to rid the metal of detrimental impurities or reduce them to within specification limits after which they are skimmed off with the aid of some added fluxes. The bath of metal is then poled with wood, chemically tested before and after additions are made to bring it up to required specifications, and finally poured into ingots, after which a final chemical and physical examination is made. As mentioned above, the impurities have either been eliminated or reduced to the point where they meet the specifications of all Government Agencies or Technical Societies. Therefore, why worry as to whether castings are to be made from strategic virgin metals or secondary ingots from available scrap. Tests have continuously been made and it is a proven fact that physical properties are no better from virgin metals than from secondary ingots of the same composition. It is definitely more fool-proof to make castings from secondary ingot than from virgin metals because if an error is made in weighing the copper, tin, lead or zinc of the latter, the analysis is off balance. If an error is made in weighing secondary ingots, you merely have more or less metal but the analysis remains constant. Furthermore, when pure copper is melted. it is susceptible to reducing gases as well as oxides. Secondary ingot is already alloyed with natural deoxidizers such as zinc, etc., and the volatilization of the latter also tends to eliminate reducing gases mechanically.

Jan 1, 1943

By B. L. Averba

THE determination of retained austenite by X-ray diffraction uses the following relationship:"" Pa = constant . RVaA (8) [1] where: P is the diffracted power from phase a; R, the calculated intensity factor; A, the absorption correction; 8, the diffraction angle; and Va, the irradiated volume of phase a For a flat sample making a glancing angle + with the incident beam the absorption correction is calculated as: * 1 sin (28 - 4) A .= 1 [2] µ sin, (28 — 4) + sin 4 where µ is the linear absorption coefficient. Beu3 has proposed that the computed absorption correction be combined with the constant and that the experimental observation of the absorption correction used in the original method be eliminated. This procedure is permissible only if: 1—There is no preferred orientation in the sample, and 2—The geometric requirements have been met precisely. If these conditions are not met the validity of the determination is questionable. Unfortunately the necessary conditions must be tested experimentally for each determination, and this is done most easily by observing whether the apparent absorption has the form of Eq. 2. In practice it may be convenient to plot P/R vs the parameter sin (2? — )[sin (28 — +) + sin +] to obtain straight lines for the ap- parent absorption correction. Deviations from straight lines indicate either that the sample has not been positioned correctly or that the relative intensities are incorrect. Geiger counter spectrometer methods also have been used successfully for retained austenite determinations. In the usual spectrometer arrangement the sample makes equal angles with the incident and diffracted beams (8 = ); thus the absorption correction is simply 1/2µ and independent of 8. The retained austenite is then easily calculated from integrated intensity measurements on austenite and martensite lines. If the austenite content is not too low and if excessive fluorescence is not encountered it is possible to use filtered radiation and a recording spectrometer. The integrated intensities are then proportional to the areas of the diffraction peaks. For lower austenite contents the following technique was used: A bent quartz monochromator and a Geiger counter spectrometer are employed. The monochromatic radiation is focused on the front slit, and the counter slit is made broad enough to include the entire line. A series of readings are taken in the vicinity of the diffraction peak. The highest reading includes the diffraction line plus background. The background is obtained from the average of several readings on each side of the peak. The integrated intensity is obtained by subtraction. CrK, and FeKa have been used for such determinations. This technique has been found to be as sensitive and as accurate as the photographic methods. In the spectrometer methods the ratio austenite/ martensite may be obtained from readings on austenite and martensite lines and the application of Eq. 1. It is also possible to use external standards and to obtain the austenite contents from measurements on the austenite lines alone. The small correction for carbides then becomes unnecessary. The best standards are probably samples with known

Jan 1, 1954

By R. C. Amero

FULLERS earth is a general name applied to claylike minerals that have high natural ad-sorptive powers. They are usually distinguished from ordinary clays by a higher content of combined moisture and a lower apparent density. The definition is based on the ancient use of the material for fulling, or cleansing woolen cloth of oil and grease; a more modern definition should mention the ability to decolorize oil and should differentiate between fullers earth, which is naturally active, and certain bentonites which only develop decolorizing power after being leached with strong mineral acids. The distinction between fullers earth and activable bentonite is not a sharp one, since all types of intermediate clays can be found that are more or less naturally active, yet more or less responsive to acid treatment. The differences seem to depend somewhat on the amount of leaching that has been effected by natural ground waters. Physical Properties The principal mineral constituent of activable bentonite is montmorillonite, a hydrous aluminum silicate to which the following formula is generally assigned: Al,Si,O,, (OH), • 12H,O Fullers earth, as well as bentonite, is usually composed of some variety of montmorillonite, but in the United States the largest tonnage of fullers earth is mined from deposits of attapulgite. Attapul- gite is a clay mineral composed of monoclinic crystals of colloidal size, having a fibrous shape similar to that of asbestos. A sectional model of the attapulgite "molecule" is shown in Fig. 1. A single unit is about 18.0 A wide and 12.9 A high. The length of the molecule is variable, but this same molecular pattern is repeated to form needlework crystals, rather than flakes, as is characteristic of the montmorillonite crystals. Attapulgite has been assigned the following formula,&apos; although actually, the magnesia has been largely replaced by alumina: (0%) . (OH), • Mg, . Si, • Olp . 4H20. The fibrous nature of the mineral is clearly shown in the electron micrograph, Fig. 2. Although the deposits contain 5 to 30 pct quartz and carbonate minerals, impurities in the commercial products are held at a low level, principally by selective mining on the basis of oil decolorizing efficiency. The particle size distribution of this fullers earth has been estimated from settling rates, using sodium citrate as a dispersing agent.&apos; Since Stokes&apos; law only applies to spherical particles, the equivalent particle diameter calculated below is probably more nearly an index of diameter rather than length of the needlelike crystals: Eqnivalent Arbitrary Particle Distribution, Size Diameter Pci Classil3catian 0.05 mm 0.80 0.80 pct sand 0.05-0.02 mm 10.96 0.02-0.01 mm 9.65 42.10 pct silt 0.01-0.005 mm 11.40 0.005-0.002 mm 10.09 2-0.5 micron 10.61 0.5-0.2 micron 29.09 56.58 pct clay 0.2-0.05 micron 14.86 0.05 micron 1.96 The differential thermal analysis shown in Fig. 3 is unique and supports the theory that this is a distinct mineral species.&apos; ~ Other general properties of the mineral as shown in Table I.

Jan 1, 1952

By R. L. Manegold

IN RECENT years there has been a decided trend toward bringing ores and coal out of open-pit and underground mines by long, sloping single-stage belt conveyors. Because the high investment cost of haulage conveyors dictates maximum speed and full loading for economy, conveyor belts are running faster than ever before, with a greater burden on the belts. In addition, larger blocks of ore are being mined and conveyed. Rougher use of large mining equipment generates more tramp iron; mechanical loading rejects nothing. Therefore the extraction of tramp iron and other deleterious materials for the protection of haulage conveyors at transfer points, chutes, and idlers is becoming an increasingly difficult job. In our experience, the best method of protecting long haulage conveyor belts is to extract the tramp iron, at least the large tramp iron, ahead of the main haulage conveyor or series of conveyors. It is most practical to collect the tramp iron at the earliest source -possible, usually directly after the primary car unloading dump or after the primary crushers, which are ordinarily not affected by tramp iron, see Fig. 1. The coal or ore from its primary source is fed by chutes or feeders onto a buffer or shock belt conveyor, a short, ruggedly built conveyor system that can be considered somewhat expendable. Its primary function is to bring the ore up to speed for transmission to the haulage conveyor. There is no assured method of protecting this conveyor from tramp iron damage at its feed end. Likewise, there is a possibility, although remote if the magnet applied to this belt is installed correctly, that the tramp iron extraction process will damage the shock belt, which is, however, short and comparatively inexpensive. Magnetic head pulleys with their maximum effective range of about 8 or 10. in. cannot be considered for large installations, even though they are desirable because of the continuous and, automatic tramp, iron removal feature. Further, the most dangerous form of tramp iron with respect to potential damage to a haulage conveyor is the long rod, roof bolt, or rail segment, which is not extracted satisfactorily because of the limited tangential contact of such an iron shape with the magnet. Magnetic head pulleys are also undesirable because the dribble or fines from the head pulley discharge at the same point as the tramp iron. Screens must be used to separate the fines from the magnetics, resulting in additional expense and loss of head room. Magnetic detectors have been used with success, but they only indicate the, presence of iron, and the conveyor belt must be stopped for its manual removal. The, stoppage of any conveyor means that all preceding equipment must be stopped or a cascade of material on the idle conveyor will result. Occasionally combinations of detectors and magnetic, head pulleys are used. The detector precedes the magnetic head pulley and detects only large pieces of tramp iron, while the magnetic pulley subsequently takes out the smaller ferrous metals. Suspended magnets overcome the disadvantages of magnetic head pulleys, detectors or a combination of both. There are several magnetic extraction problems involved in applying a suspended magnet at this

Jan 1, 1952