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By L. Rodd, C. Walker, T. Koehler

"The energy content of slag represents approximately 80% of the total energy inputs to a laterite nickel electric smelting furnace. Conventional slag handling methods produce only low grade heat and the entire thermal content of the slag is wasted. A heat recovery system proven on BOF slag in the mid 1980's can be adapted to a laterite nickel rotary kiln electric furnace (RKEF) plant. Slag is granulated using a high velocity air jet inside a waste heat boiler to produce slag granules. Granules are cooled during flight via radiation to water-cooled boiler walls and by convection to the granulation air. This system has the potential to recover 80% of the total slag energy in the form of hot air and steam.Integration of the recovered slag energy into the RKEF smelter can be achieved by a number of methods, including the use of preheated air at the ore dryers, kiln or existing power plant. Energy from the slag boiler may be used to reheat boiler feed water from the power plant or to produce superheated steam and generate electricity. The heat recovery system can provide up to 25 - 30% of the total electrical requirement of each RKEF line and a reduction in CO2 emissions of 234,000 tonnes per year.The total cost of slag heat recovery, including CAPEX and OPEX, is $2.5/GJ, which is competitive with using coal when CO2 credits are $18 per tonne. With Clean Development Mechanism (CDM) carbon credits currently selling at €13 ($20 USD) per tonne CO2, there is considerable opportunity for slag producers in regions including South America, Korea, Indonesia and Africa. Increasing fuel costs and more wide-spread legislation for GHG emissions will likely create conditions in which slag heat recovery could generate a favourable return on investment."

Jan 1, 2010

By H. D. Kessler, Joseph B. McAndrew

WHEN there is occasion to make structural use of metals at temperatures above 900°C (1652°F), the choice of alloys is severely limited, and those materials which meet special requirements as to density, strength, toughness, and resistance to creep, oxidation, or fatigue at such temperatures are at a premium. Any new type of high temperature alloy which is found to be outstanding with respect to several of these properties is therefore likely to find special applications for which it is uniquely advantageous as compared to other alloys. Such is the case with the 7 titanium-aluminum alloys. This y phase is an intermediate phase based on the composition TiAl, and extends from about 35.5 to 44.5 wt pct Al at temperatures up to 1000°C, above which the field broadens to include 60 pct A1 at 1340°C.' The alloys of principal interest have a narrow melting range, with 7 forming by the peri-tectic reaction, melt + ß=y at about 1460°C. Unlike most intermediate phases, the binary alloys containing titanium and aluminum in approximately equiatomic proportion possess low room temperature hardness,' in addition to which they have low density, good resistance to oxidation, and relatively high modulus of elasticity.' Unpublished work done at Armour Research Foundation showed

Jan 1, 1957

By Thomas Jancic

Two of the open-pit, copper-molybdenum mining operations of Duval Corporation are located in the foothills of the Sierrita Mountains approximately 30 miles southwest of Tucson, Arizona. In March 1959, mining from a moderate sized porphyry deposit known as the "Esperanza orebody" was begun at a rate of 10,000 tons of ore per day. Some five years later, a sizeable group of mineral claims west of Esperanza became available to Duval. Earlier exploration and development drilling in this area by others indicated an orebody of greater size but lower tenor than that of Esperanza. A drilling program conducted by Duval not only verified the results of the prior drilling but in addition, revealed a continuation of the mineralization to the south and east. Thus was developed the Sierrita orebody. Open pit design delineated 414 million tons of ore containing an average copper content of 0.35 percent and a molybdenum content of 0.036 percent. In April 1966, the office of Emergency Planning authorized General Services Administration to undertake a program under the Defense Production Act to en¬courage additional domestic production of copper in the interest of national defense. Duval's Sierrita Project was brought to GSA's attention in May 1966 on November 28, 1967, an agreement providing for the development of the Sierrita Property was consummated with GSA. In March 1970, the first copper-molybdenum concentrates were produced in the 72,000 tons-per-day concentrator. Pre-production stripping exceeded the scheduled 105 million tons despite the comparatively short lead time available to equip the mine, build the mine plant facilities, and staff the technical, supervisory, and labor forces. Shortly before the first concentrates were produced, plans were being formulated for a program to increase processing capacity by approximately 15 percent, which entailed increasing overall mining capacity by approximately 50,000 tons per day. This program is scheduled for completion early in 1971.

Jan 1, 1971

By Terry I. Mudder

Metal complexed cyanides in wastewaters form as a result of interactions of free cyanide with metals present in the wastewater and exhibit varying degrees of stability, toxicity, and treatability. Thiocyanate, a pollutant commonly found in cyanide bearing wastewaters, is formed through the interaction of free cyanide with sulfur containing species (i. e. pyrrhotite) both present in the wastewater. In certain industrial processes, such as the beneficiation of gold and silver, cyanide is an essential reagent. Since free cyanide, complexed cyanides, and thiocyanate are potentially toxic to humans and aquatic organisms, these compounds must be removed from wastewaters prior to their discharge into surface or ground waters serving as potential potable water sources, marine or fresh water habitats

Jan 1, 1984

By K. S. Liddell

Ultra fine grinding is becoming increasingly used in the mining industry over recent years. The development of large-scale units for high volume grinding of low value materials and concentrates has opened up new areas of economical metal extraction. The ability of stirred vertical mills to produce very fine particles with high surface areas, is providing mine operators with novel extraction possibilities within gold concentrate regrind for Biox or leach feed, Activox autoclave feed preparation and lime slaking operations. The paper will review design, operational and maintenance aspects of ANI-Metprotech stirred vertical mills, comparing these to tower mills, horizontal bead mills and attrition mills.

Jan 1, 2001

By P. Eger

Minnesota contains a major low- grade copper-nickel sulfide resource within the Duluth Gabbro Complex. Several companies have actively explored this resource. Open pit mining would be required to develop portions of the resource, and as a result, large quantities of sulfide-bearing lean ore and waste rock would be stockpiled. Although the problems associated with high-sulfide tailings and acid mine drainage are well documented, the leaching characteristics of low-grade mineralized stockpiles must be understood for environmental impact assessment, siting, and planning. The Minnesota Department of Natural Resources and AMAX have cooperated in a research program to study leaching and revegetation -of test stock- piles, Six small (1600 'tons) stockpiles were constructed at the MINNAMAX Project Site and they range in composition from 0.29 to 043 g % copper, 0.075 to 0.09% nickel, and 0.6 to 1.45% sulfur. Two of the stockpiles were covered with mineral soil, one was covered with topsoil, and three were not treated Revegetation plots were established on the piles which received a soil cover. Leachate samples and flow measurements were collected at the base of the piles. Initial results indicate that metal concentrations in leachate samples range from 10 to 10,000 times the concentrations found in the streams of the area. During the first growing season, the test pile covered with topsoil yielded 30% less water than untreated test piles. No reduction in yield was observed for the piles that were covered with the less fertile mineral soil. First-year vegetative growth, expressed as biomass, was four times greater on the-topsoil pile than on the mineral soil piles. The long-term effects of vegetation on leachate quantity and quality continue to be investigated.

Jan 1, 1980

By S. Maurhofer

The Uetliberg Tunnel in Switzerland is the longest tunnel in Zurich?s new Western Bypass Expressway, connecting the Birmensdorf bypass in the west with the existing Zurich-Chur national motorway in the east. Thus the tunnel links the Zurich-West motorway interchange with the Zurich-South interchange (shown in Figure 1). The Uetliberg Tunnel project comprises two parallel tubes, each about 4.4 km long. The two tunnel tubes are connected every 300 m by a transverse walkway and every 900 m by a transverse roadway (Figure 2). The SOS niches are spaced 150 m apart. Portal stations with machinery rooms are located at the west and east portals. The tunnel falls with a gradient of 1.6% from the west portal at Wannenboden to the east portal at Gänziloo. The Project costs are about CHF 1.12 billion. The carcass construction started in March 2001. The Opening of the Uetliberg Tunnel is planed in spring 2009.

By D. B. Gawthrop, A. Taylor, G. St. J. Ferrott

"This paper presents data obtained as part of an investigation of the explosive properties of ammonium nitrate. During the course of the work: it was found desirable to investigate the sensitizing action of nitrostarch in mixtures of ammonium nitrate containing low percentages of this material. Data were derived as to the sensitivity to decomposition by heat and the completeness of detonation and,of detonation of straight ammonium nitrate, as compared with ammonium nitrate-nitrostrarch mixtures containing 1 to 10 per cent of nitrostarch.The ammonium nitrate was a sample of coated material obtained from the salvage of amatol shells. Its analysis was as follows:"

Feb 1, 1930

By R. E. Siemens

The Bureau of Mines sought to correlate the variation of structure-sensitive superconducting properties with microstructure in selected Cb-Hf alloys and to compare the experimental results with those predicted by the GLAG theory. Alloys in wire and rod form were evaluated by critical temperature, critical current, magnetization, resistivity measurements, and optical and electron microscopy. In the absence of an external magnetic field, the current carrying capacity of columbium was greatly reduced with addition of up to 85 wt pct Hf, but a large increase in Ic occurred for Cb-90 wt pct Hf. Favorable Ic-H behavior was observed in cold-drawn alloys containing 25 and 57 wt pct Hf. Correspondingly, a maximum in Tc occurred at 10-25 and 90 wt pct Hf. Four- or five-hour heat treatment at 600° C resulted in maximum Ic values and maximum hysteresis in magnetization. Dislocation pileup, polygonization, and fine precipitates, responsible for flux pinning, are listed in order of their effectiveness as pinning sites. Sites having an average separation of 1,500 to 3,000 A and a diameter of 1,000 A or less were most effective in influencing Ic-H and M-H behavior. The resistive critical fields increased smoothly to a maximum of 80 k0e between 60 and 65 wt pct Hf. Highfield, mixed state, reversible paramagnetism was observed in some alloys; however, the GLAG theory was in good agreement with experimental results for the majority of the composition range studied.

Jan 1, 1969

By G. Barr

"The Polymetallic Twin Metals Project deposits in Northern Minnesota are being developed by Twin Metals Minnesota. Amongst the flowsheet alternatives being examined at a prefeasibility study level is one involving flotation of a bulk concentrate followed by high temperature autoclaving of the concentrate using the PlatsolTM Process to solubilize all the valuable metals in the concentrate (Cu, Ni, Pt, Pd and Au). Autoclave operating conditions have been established in which recoveries of all the pay metals are in the 92 to 99% range, and the autoclave residence time has been reduced to an hour. The stability of the precious metals in solution after autoclaving has been assessed, as well as the impact of internally recycling process liquors to the feed to the autoclave.INTRODUCTIONPlatinum group metals (PGMs; platinum, palladium, rhodium, ruthenium, iridium and osmium) are traditionally recovered from their ores by production of a high-grade concentrate by flotation and/or gravity concentration followed by smelting of the concentrate. Concentrates are smelted to remove nonsulphide minerals and converted to remove iron and sulphur. The resulting high PGM grade matte is processed hydrometallurgically to leach the base metals, leaving a residue with a very high grade (20-40%) PGM concentrate for final refining to individual PGMs. Due to the complexity of this process and its high capital cost, this process can only be applied economically to high grade PGM ores. Lower grade PGM concentrates might be accepted as copper and nickel concentrates but may not yield attractive payment terms for the PGMs. The mineralized Duluth Gabbro region of Minnesota hosts several low grade base and PGM deposits and Twin Metals Minnesota is working to develop the Twin Metals Project deposits. Earlier studies at both pre-feasibility and feasibility level, by other mining companies, have failed to yield an economically viable project based on the traditional process. It is possible to produce a saleable copper concentrate from these ores, but economic recovery of nickel and PGMs into saleable concentrates proved elusive.In the late 1990’s and early 2000’s the Platsol™ process was developed at SGS in Lakefield, Ontario (at the time Lakefield Research Limited) for the NorthMet deposit (also within the Duluth Gabbro) owned by PolyMet Mining. The process developed was a high temperature chloride-assisted pressure oxidation process capable of dissolving simultaneously the base metals (Cu, Ni, Co, Zn), gold and the PGMs."

Jan 1, 2012

Trial ore production began at the Hilton Mine in 1987 with ore being trucked into Mount Isa for treatment at the Lead/zinc Concentrator. The expansion of output from the Hilton mine required additional concentrator capacity which led to the design and construction of the Hilton Concentrator. Systematic laboratory and Pilot Plant testwork dating back to 1968 to develop a flowsheet and reagent regime suited to the varied and generally complex mineralogy of the Hilton ore was successfully concluded with a Pilot Plant trial using a sulphoxy based sphalerite depression system in early 1988.In May 1988, a decision was made to build a 750 000 t/y concentrator with some novel technological and capital-efficient features. The Hilton Concentrator commenced operations in December 1989 reaching target grinding rates and metallurgical performance by mid-1990. Current ore treatment rate is 800 000 t/y.Lead concentrates assaying 43 - 48 per cent Pb are combined with the Lead/Zinc Concentrator lead concentrates and smelted at Mount Isa. Zinc and Low Grade Middlings (LGM) concentrates are filtered in the zinc filter plant at Isa with those from the Lead/Zinc Concentrator and then railed to Townsville for export or treatment in Australia.

Jan 1, 1993

Study on Treatment of Wastewater Containing Heavy Metal Ions Using Modified Oceanic Cobalt-Rich Crusts

Sep 13, 2010

By T. A. Rickard

FROM AN OLD MANUSCRIPT THE text for this contribution to the history of the Patio process is a page taken from a Spanish manuscript of Bartholomé Arzaj Sanches y Uela, now in the possession of the library of Brown University (see Figure 1). I owe the photographic print to the courtesy of the librarian. The date, as will be seen, is 1577. The subject is the treatment of silver ores at Potosi, in old Peru, but in modern Bolivia. The group of drawings (see Figure 1) is entitled "Arrangement of the machines in which the ores of silver are ground on the bank of Potosi, the one of two batteries and the other of one". First, top-left, we have the sketch of a "machine of one battery" operated by a water-wheel. From this mill the water runs in part along an aqueduct supported by pillars to the two-battery mill and in part down the hillside, where men are at work with "tubs, which are called cochas, in which are washed the charges of ore and amalgam". Along the top of the page are shown the "various offices in which are stored the copper, the salt, the ucha [llama dung, used for fuel], the lime, and other materials for the re-duction of the ores". To the right of these are "apartments and warehouses in which is stored the ore brought down from the mountain"?the cerro de Potosi. Through the arched portal come the "sheep [llamas] on which is brought down the ore". To this is added a reference to "the second gate, which closes the whole works". On the level above the aqueduct are seen six men carrying skin-bags that contain water. This is being sprinkled on the ore. On the extreme right are six more of these water-carriers. Next comes the two-battery mill. The mortero, or mortar, is marked on the right. Another description says, "Stone, on which the stamps crush the ore, which is called a mortar". To the left of these are indicated "screens of wire in which the ore is sifted after it has been crushed". Farther to the left, near the ore-washers, is a "shed, in which is placed the metallurgist to test the ore of the boxes". Along the left side of the page are sundry buildings, one of which is labelled a "shed in which are placed the cones of silver". Near it, lower, to the right, is the "furnace, in which are roasted the sulphides to facilitate treatment". Farther to the right is the "place where the Indians mix the charges of sulphide ore, which is called a buitron". Alongside is a "buitron in which the charges are prepared for reduction". Inside is a container, to which are attached the words "measure, called topo" On the extreme right is "another buitron in which the Indians mix the ores with their feet, as well as the heaps for the preparation of the reduction". Near the bottom, on the extreme left, is the "chapel where the holy sacri¬fice of the mass is celebrated". Under the first buitron is the "place where they dry the moist ores, called pãpear".

Jan 1, 1935

By R. Abela

In over ninety per cent of mineral sand plants, the most commonly used equipment is the spiral concentrator. The tonnage treated over typical spiral plants can range from as low as 200 ton per hour up to 6 000 tons per hour. Due to the limited capacity of each spiral, plants consist of multiple spirals assembled into banks. The mixture of minerals present in the sand feed requires several spiral stages to achieve the desired mineral recovery. Due to the complexity of the spiral circuits, there are always some design and operating problems that present themselves in plants. These range from slurry distribution, spirals operating at incorrect design parameters and feed variations. The material being treated is difficult to maintain in suspension due to its high density, resulting in segregation and preferential flow in pipelines, distributors and sumps. The objective of this paper is to highlight common problems associated with spiral installations to designers and operators of plants. Practical solutions are recommended in order to prevent and minimize these mistakes in the design. These solutions may be used to improve performance on operating plants.

Jan 1, 2003

By Seth T. Reese

THE INJURY DATA and experience at coal mines in the United States for 1949 are presented in this bulletin under the following general sections: (A) General injury experience: 1. Tables 1 to 18: Contain current data on number of injuries, frequency and severity rates, average severity, major and detailed classification of injuries, and distribution of injuries by part of body and days of disability. (B) Selected injury data: 1. Tables 19 to 45: Contain current data on various selected factors of operating and physical conditions that have contributed to the occurrence of injuries at bituminous-coal mines and at Pennsylvania anthracite mines. These factors include extent of hand and machine loading, size of mine by employment and production groups, thickness of bed or vein, duration and number of work shifts, rating of mine as gassy, or nongassy, types of cap lamps, ventilation, explosives, and combinations of factors that influence mine safety. (C) Injury experience by States: 1. Tables 46 to 52: Contain current data on number of injuries, frequency and severity rates, average severity, major classification of injuries, injury records of all coal-producing States, and the relative standing by States according to injury-frequency rates. (D) Appendix: 1. Tables 53 to 84: Contain certain injury data for the years 1943-47 that have not heretofore been published.

Jan 1, 1953

By E. C. Strycula

Solid-state equipment offers a number of potential advantages in the coal mining industry. These advantages can only be realized, however, through careful and judicious design of the control and protection circuitry. Problems relating to the interaction of solid-state equipment on the electrical system and the special design considerations which apply to solid-state circuitry are examined. Design guidelines are provided for solid-state applications in the areas of: (a) equipment operated from a DC trolley wire, (b) drive systems for surface mining equipment, (c) AC motor starters, and (d) electromagnetic interference.

Jan 1, 1981

By Edwin Ludlow

EDWIN LUDLOW, the 41st President of the A. I. M. E., died in Muskogee, Okla., on Feb. 10, 1924, after a brief illness of influenza followed by pneumonia. He was born in Oakdale, Long Island, N. Y., March 12, 1858, a son of William Handy and Louise (Nicoll) Ludlow, attended Flushing (L. I.) Institute 1868-75: and Columbia School of Mines (N. Y.) where he graduated with the degree of Engineer of Mines with the Class of 1879. After graduation his first employment was as assistant engineer in river and harbor work on the Delaware River near Philadelphia. In 1881 he went to Mexico as one of the engineers on the Mexican Central Railroad, then building from Laredo, Texas, to Mexico City. The road being in financial difficulties he returned to the United, [states, and shortly thereafter was made superintendent of the Union Coal Co.'s collieries located in the western middle anthracite field, between Shamokin and Mt. Carmel. Ludlow and Arthur H. Storrs, then superintendent of the Neilson Shaft Colliery, in Shamokin, were known throughout that section as the baby superintendents, both under 25 years of age. The appointments of such young men were at that time unique and set a record for youthful assumption of re-

Jan 1, 1924

By P. D. Laverty

This report describes a processing method developed by the U.S. Bureau of Mines to produce high-quality flake graphite from the steelmaking waste known as kish. The kish produced by current steelmaking practices is a mixture of graphite, desulfurization slag, and iron that is skimmed from the molten iron feed to the basic oxygen furnace. It is estimated that the graphite content of kish discarded by U.S. steel plants is more than sufficient to meet the total U.S. demand for flake graphite. That need is now filled by natural graphite from foreign sources. Kish was treated by a combination of screening and hydraulic classification to produce a concentrate containing greater than 70 pct graphite. Leaching of the concentrate with hydrochloric acid solution gave a graphite product with 95-pct purity. An optional secondary leaching operation with hydrofluoric acid produced graphite with a purity of 98 pet or greater. The flake size of graphite from kish ranged from 10 mesh down. Evaluation of test samples by industrial graphite users indicated that kish graphite is a suitable substitute for the natural material for most uses. A process flowsheet and material balance for pilot plant design are presented.

Jan 1, 2010

By L. F. Pett

ALTHOUGH Kennecott's orebody has long been outlined, it is still necessary to define further its limits. This mine, long an advocate of churn drill methods, recently supplemented its practice by using a diamond drill. So far, the diamond drill holes have been along the main canyon floor which traverses the mine area. This zone, highly fractured, is cut by numerous friable quartz veinlets encountered within the porphyry. These veinlets, formed by filling contracted porphyry fractures with hot, siliceous, mineral-bearing solutions, are often more highly mineralized than adjacent porphyry. They pulverize on contact, yielding only slight amounts of core. For this reason neither analysis of core recovered (amounting to 37.5 pct of approximately 10,000 ft of hole drilled) nor sludge analysis considered separately would give correct assays. The following equation has been adopted: actual weight theoretical actual weight core — theoretical weight sludge Drilling is done un.der contract. Collection of core and sludge samples together with double-checked assaying is done by the company. Microscopic, specto-graphic, and X-ray examination of rock specimen saved is carried on at the University of Utah. A fund, set up by Kennecott Copper Corp. at the University, has encouraged cooperation in solution of mining research problems. The drill used is a Longyear hydraulic powered by a Waukesha 26-hp gasoline motor. A pump, Bean Royal type, is powered by a 7-hp Briggs-Stratton gasoline motor. Weight of unit is 4000 lb and is placed under tripod of 35-ft wood poles within a 25-ft sq area. The following results have been attained: Avg depth of hole drilled, ft 1215 Avg drilling depth per 8-hr shift, ft 10.11 Core Size, In. Avg depth of NC casing, ft 21 2 11/16 Avg depth of NX casing and size core, ft 178 2Ye Avg depth of BX casing and size core. ft 427 1% Avg depth of AX casing and size core, ft 585 1 5/32 Through the use of metal sludge collection boxes, 12 ft long, 18 in. wide, and varying from 12 to 24 in. in depth, sludge recovery has been maintained above 95 pct. Samples, both core and sludge, are assembled for each sample interval and averaged in 50-ft bench increments. Determinations for mineral values are made at both the mine and mill. Assay results varying in excess of 0.05 of copper assay values are rerun to determine correct assay. The organization chart for the Kennecott Utah copper mine is shown in Fig. 1. The Utah mine requires approximately four men per shovel shift to drill solid rock, properly slope banks, and trim loose material from the top of bench sections in making them safe for men and equipment working below. During recent years, while the mine has been worked day and night, powder crews have worked day shift only. Series of toe holes drilled back of advancing electric shovels comprise approximately 70 pct of all mine drilling. These holes, spaced about 21 ft apart, have a depth of 24 ft. Of the footage drilled, 25 pct is for vertical holes either 24 or 28 ft deep, spaced about 28 ft centers along the top rim of banks. This combination of toe holes, using an Ingersol-Rand DA 35 drifter mounted on a tripod and the vertical holes using an Ingersol-Rand Type X71 wagon drill is standard practice above the pit elevation. These benches are 68 ft high, with a few exceptions, while benches below the pit elevation are 50 ft high and require very little drilling other than toe holes. Five benches, varying in height from 80 to 100 ft, require a combination of toe holes plus vertical holes and bank holes or electric churn drill holes. Bank holes using the DA 35 drifter and electric churn drill using Bucyrus-Erie Model 29T, account for 5 pct of feet drilled. Until the past year, drilling of fill material on upper levels was accomplished by churn drill method. Wagon drills are now used for fills because they can be set up and drill a hole an hour. It is necessary to run an air drill 7 hr for each 8-hr shovel shift. Average drilling rates during the past year were as follows: DA 35 drifter using 2 and 3-man crews — 4.36 ft per man hr X 71 wagon drill using 2-man crews — 6.45 ft per man hr DA 35 drifter bank holes, 3-man crews — 2.53 ft per man hr Bucyrus-Erie churn drill, 2-man crews —1.9 ft per man hr Total mine avg — 4.57 ft per man hr Mine practice assigns powdermen to job locations where they perform all drilling and blasting work within the area. Almost as much labor is consumed in springing and loading holes as in drilling them. Average powder consumption is approximately as follows: Per Electric Shovel Shift, Lb To spring and chamber holes — 50 Blasting — 500 Trim banks — 50 One fifth of all powder used is Hercules hercomite bag powder while four fifths is Hercules gelamite 65 pct strength in 1 Vi in. x 8 in. sticks. Frozen material, overhanging from top banks, creates a winter hazard to shovel operators and equipment. Experiments are being conducted with a high-speed Hercules powder in an effort to eliminate this condition. Banks are trimmed by powdermen who lodge 8 or 10 bundles of powder, containing 15 or 20 sticks each, across the upper surfaces of each working face. To each bundle of powder is attached a primer with 6-ft fuse which is lighted as the powdermen, who have lowered themselves down the bank on ropes, ascend to safety. Three angle dozers, D8 caterpillars, make grades and move supplies for the average 31 drilling machines working daily. Compressed air is furnished by three Ingersol-Rand compressors, each with rated capacity of 3750 cu ft of air per min at 100 lb reservoir pressure.

Jan 1, 1952

By J. E. Berg

This paper3 describing the mining methods at the Page mine operated by the Federal Mining & Smelting Co., is one of a series being prepared by the United States Bureau of Mines on mining practices, methods, and costs in the various mining districts of the United States. The Page mine is the most westerly of the producing lead-silver-zinc mines in the Coeur d'Alene mining district of Idaho, being situated 5 miles west of the city of Kellogg and at an elevation of 2,677 feet above sea level. It may be classed as one of the new active producers, of the district and owes its present activity largely to the development of floatation.

Jan 1, 1930