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By John Jermain Porter

(San Francisco Meeting, October, 1911.) THE value of any particular ore, coke, or limestone, for iron-making, depends upon its effect, first, upon the quality or value of the resultant product; and second, upon the cost of smelting. The factors under the first head are the percentages of phosphorus, manganese, and other elements, which remain in the pig-iron ; while those under the second head are the percentages of iron, carbon, or lime, the slag-forming constituents, which must be fluxed, and any other ingredients which give rise to variations in cost of fuel and flux. The first of these criteria of value is perhaps not capable of being reduced to a general formula embracing all cases, since it must vary to a great extent with market-fluctuations. For example, Bessemer pig is worth sometimes more and sometimes less than foundry-pig ; and the premium paid for high manganese varies within wide limits. The second, or smelting-value, however, can be expressed numerically with a fair degree of accuracy; and it is with this phase of valuation that the present paper is concerned. The utility of a means of comparing numerically the relative smelting-values of different ores, cokes, and limestones, is self-evident, although it may not have importance in Northern practice, where materials are largely standardized. On the other hand, there are, especially in the South, many furnace-managers who must use a great variety of materials of widely varying grade; and it is with special reference to this district, and in connection with the efficiency-methods which I have elsewhere proposed,1 that the methods herein described were devised.

Mar 1, 1912

By H. Kenyon Burch

INTRODUCTION THE Inspiration Consolidated Copper Co.'s plant at Miami, Ariz., was designed and built to make possible the profitable working of a low-grade, finely disseminated copper deposit containing 100,000,000 tons of ore averaging 1.64 per cent. in copper. From the-beginning it was evident that the-plant could not, be kept integral but that a break would have to be made somewhere in the flow-sheet, removing at least the concentrator to a site more suitable than any available near the mine. It was finally decided, after considering numerous arrangements, to do the coarse crushing at the mine, to store the crushed ore in a bin from which it could be loaded into railroad cars and to haul it to the concentrator, an excellent site for which was available about 1 ¾ miles from the mine. The original intention was to equip a plant to treat 7,500 tons of ore per clay, but through the acquisition and proving up of additional ore reserves, the introduction of the Ohio caving system, and the excellent results obtained in the test mill (which made it possible to treat a lower-grade ore than had been thought possible) it was evident that a plant of much greater capacity should be supplied. It was, therefore, decided to treat approximately 15,000 tons of ore per day, the duties of the four main divisions to be as follows:

Jan 9, 1916

By Robert M. Abrams

The passage of an ore pulp through the froth flotation process results in the production, after dewatering, of a finely divided concentrate. This intermediate product must be handled to the subsequent process facilities via beit or other types of conveyors, hoppers, bins, chutes and feeders, and the like. Flotation concentrate has some difficult handling properties such as abrasiveness, sluggishness when moist, or fluidity when dry. These characteristics require the judicious selection or design of handling equipment and the careful engineering of transfer chutes. It is to the special problems which are encountered in this in-plant handling procedure that this paper is addressed. The methods which will be discussed have been developed primarily for the handling of the froth flotation concentrates of copper, lead, and taconite iron ores, but may be generalized, with appropriate modifications, to apply to other similar metallic concentrates. The controlled feeding of concentrates from bins and hoppers is covered along with some of the techniques used to transfer concentrates from one belt conveyor to another. Methods are described for unloading belt conveyors into bins or storage piles. The reclamation from storage piles is also treated.

Jan 1, 1979

By L. W. Casteel, E. A. Jones

MINING the lead deposits of Southeast east Missouri has reached a high stage of technical development dictated by the scattered occurrences of low-grade ore through favorable horizons in the Bonne Terre formation. The rock-breaking, handling, and related practices described here have evolved as the best way of mining these particular deposits. These methods are applied in the production of about 19,000 tons of lead ore per operating day from four mines in Southeast Missouri. Development drilling Drifts 10 ft wide and either 8 or 9 ft high are driven in this district on contract. Miners are guaranteed standard wages plus a bonus on the footage advanced, if earned. Labor, fuse, caps, wire, and dynamite used are charged against the contract. Normally the cycle of operations in drifting is performed by a two-man crew, a motor operator and a shovel operator who first load out the rock broken by the preceding shift. The drift crew then sets up two vertical column and mounts the drills. Cleveland HC-10's are used which have a steel lug integral with the cylinder for connection with the air feed. All drilling in drifts is wet. Air pressure is 85 to 100 lb at the face. From 28 to 32 holes are drilled per swing in a pyramid cut. Cut holes

Jan 1, 1947

By Gregory L. Gullord

In September, 1975, production was begun at the new milling facility of ASARCO Incorporated, the Young Mill. The mill is located adjacent to the Young Mine 35 km (22 miles) east of Knoxville, Tennessee. The Young Mill replaced the obsolete Mascot Mill which had been operating since 1913. The basic flow sheet of Mascot was retained for the construction of the Young Mill, with a 20% increase in capacity included in the design. The Young Mill is capable of processing 7710 - t/d (8500 stpd) of crude ore from three underground mines (Young, Immel and Coy) yielding 300 - t/d (330 stpd) of 63% zinc concentrate. The Young Mill is unique in that 92.5% of the ore milled is recovered as saleable products, including zinc concentrate, construction aggregate, sand and agricultural limestone. Processing at the Young Mill is divided into five basic stages: two-stage crushing, heavy media separation, grinding-flotation, concentrate filtering and drying, and limestone by-products recovery. The ore bodies are located in the Knox formation, which consists of primary dolomite, secondary dolomite and limestone with silicas in the form of chert and jasperoid. The zinc mineralization occurs as an exceptionally pure sphalerite. The ore contains an occasional abundance of pyrite and traces of fluorite and barite. TWO STAGE CRUSHING Ore enters the crushing section from two points. The Young Mine ore is hoisted into a 816 ton (900 st) bin which feeds a 1.52m (60 in) conveyor via a 1.52 m (60 in) vibrating feeder. The Immel ore is trucked 10 km (6 miles) by an independent contractor and weighed on automatic recording truck scales at the entrance to the property.

Jan 1, 1977

By Carl F. Floe, Michael B. Bever

ACCELERATED by the demands of war, research and development work in nonferrous physical metallurgy has continued at a rapid pace during the past year. In particular, advances have been made in processing and fabrication, especially of the light alloys, and in the surface treatment of metals. Theoretical physical metallurgy, although less active, has not been neglected and the year has brought forth some -notable contributions to our understanding of the fundamental structure of metals and their behavior in service. Of course, much work that has been done cannot be publicly discussed because of wartime restrictions.

Jan 1, 1945

Copper ores from small mines in the area south of Tucson, where the Twin Buttes mine is located, were first operated in the 1880's. Today this area is one of the great copper mining centers of the world. In 1963, Anaconda Company acquired the property from Banner Mining Company and in recent years joined with AMAX in a 50/50 partnership to work the Twin Buttes property. From the beginning, Anaconda had relied on conveyors to remove overburden, oxide material for eventual leaching, and ore from the open pit. In the large st preproduction stripping program in mining history, Anaconda in a four-year period (1965 to 1969) removed 266 million tons of overburden of an average depth of 500 ft (152 m) lying over the ore body. The alluvium overburden was stripped by scrapers carrying 70 to 80 tons and push-loaded by two D9 tractors operating in tandem. The scrapers delivered their loads to bins which fed a 60 in. (1.52 m) conveyor having a capacity of 8500 tph. The 25% slope belt conveyor, which was 2100 ft (640 m) long, transported the material to the brow of the pit, where it was transferred to an 8300 ft (2530 m) surface belt to a 1000 ton loadout bin. At this point, the overburden was trucked to the sites of the tailings ponds, where it was used for dam construction. During 1976, the one billionth ton of overburden was removed from the Twin Buttes mine. The Twin Buttes quartz monzonite porphyry contains only scattered mineralization of which chalcopyrite is the most abundant mineral, associated with chalcocite, covellite, minor bornite, and several oxide minerals. Oxidation has progressed to variable depths of up to several hundred feet below the base of the alluvium. Because of the depth of the sulfide and oxide ores and hard rock waste, two in-pit crushers and conveyor systems were also installed for their transport out of the pit; and in 1976 a third crusher and conveyor system was installed lower in the pit for waste rock only. The two initial crushing and conveying systems were designed to handle any of three functions: crushing waste to be disposed of by a rail-mounted stacker -conveyor, crushing oxide ore for conveyor and then truck transport to stockpiles for future leaching, and primary crushing of sulfide ore for conveyor transport to the two 30 in. by 70 in. (0.76 by 1.78 m) secondary crushers and the two 7 ft (2. 13 m) standards and four 7 ft shortheads in the third and fourth stage crushing plant. A third standard crusher has been in- stalled recently. Mining is done by 15 yd (11.5 cum) shovels which load 100- and 150-ton

Jan 1, 1978

R. Trimble, Pittsburg, Pa (communication to the Secretary) : There are in the proposed specifications only two points on which I wish to comment at this time.

Jan 1, 1902

By Joseph C. Buchen

IN principle, production of salt from sea water is a simple operation. Sea water is trapped in ponds, the sun and wind cause evaporation of the water, and what is left is principally salt. Commercial production of solar salt, however, requires a careful consideration of many factors. Obviously, sufficient evaporation must take place throughout the making season to produce an economically adequate crop. Other factors are density of soil, contour of ground, height of ground above sea level, minimum dilution from fresh-water streams, and distance from market. Unfortunately, in the United States the localities of greatest evaporation, allowing for the other factors, are not the most suitable for the manufacture of sea salt by solar evaporation. The south end of San Francisco Bay most

Jan 1, 1937

By AUGUSTUS J. MONKS, Norman L. Weiss

THE Santa Barbara Unit of the Compania Minera Asarco, of which the San Diego mill is a part, is in the Parral District of southern Chihuahua. Although the concentration of sulfide ores has been practised in Santa Barbara since 1900, the San Diego mill, completed in 1921, was the first plant in northern Mexico to attempt the treatment of oxidized lead ores. From the milling point of view the many ores that have been treated at San Diego are very similar, the differences among them being rather of degree than type. Since 1921 the number of ores from different sources treated simultaneously has varied from four to eleven; in no case have any distinctive characteristics warranted the separate treatment of any of these. The zone of oxidation, source of the ores milled at San Diego, extends to a depth of 300 to 400 ft. below the surface. These ores carry more gold and less silver, lead, copper, and zinc than the underlying sulfide ores; the gangue minerals appear comparatively unchanged by oxidation and leaching. Lead occurs in diverse minerals, chiefly in cerussite, galena, and angle-site, the proportion of the lead in galena varying considerably in the different mines. An average of the feed to the mill would show the following approximate distribution of the lead content:

Jan 1, 1930

By Ray W. Carpenter, Ralph Hultgren, Raymond L. Orr

Heats of formation of Au-Zn alloys of compositions Au3Zn and AuZn were rneasured at several temperatures by liquid tin solution calorimetry. The data for Au3Zn show that much smaller heat and entropy effects are associated with the a2- a1 transformation than for the a1 - a transformation. This result is consistent with reported X-ray diffraction studies which indicate that the a1 and (a2 phases have ordered structures which only slightly differ, whereas the a phase is disordered. The heat of formation of AuZn(ß') does not change significantly between 322o and 8000 K, confirming that the ß' alloy remains highly ordered to high temperatures. Not many measurements have been made of the energies of order-disorder transformations in alloy phases. In many cases these energies may be conveniently determined by measuring heats of formation as a function of temperature. Many alloy phases approximately obey Kopp's law of additivity of heat capacities; as a consequence their heats of formation do not change with temperature. In ordered alloys, however, a decrease in degree of order with increasing temperature will contribute an endothermic component to the heat of formation. Thus, for ordered phases where other anomalous contributions to heat capacity in the alloy or its pure components are absent, a measured change in the heat of formation may be attributed to the energy of disordering. A portion of the Au-Zn phase diagram1" is shown in Fig. 1. The composition AuZn(ß') has a well-ordered superlattice structure of the CsCl type at room temperature. The alloy is considered to remain ordered to its melting point,1 although, from X-ray diffraction data and electromotive force measurements of zinc activities, Terpilowski3 concluded that considerable disordering began at -673°K. The Au3Zn composition is a disordered fee solid solution (a) at high temperatures. Between 693o and 543oK, the ailoy orders to form the a1 phase, having a structure based on the L12(AuCu3) type, but with a regular distribution of ordering faults in the c direction, making it tetragonal. Below 543oK, the alloy orders further by means of slight atomic shifts, without diffusion, yielding a complex long-period orthorhom-bic superlattice structure, a2.4-7 Heats of formation of Au3Zn(a2) and AuZn(ß') were measured at 363°K by Biltz et al.8 by aqueous solution calorimetry. Heat capacities of several alloys between 14.9 and 38.4 at. pet Zn in the range from 373" to 723°K were measured by Iwasaki ef al.6 Stoichiometric Au3Zn, however, was not included in the alloys measured, the nearest composition being 26.1 at, pet Zn. Also, the heating rate used for the measurements, 2 deg per min, may not have permitted equilibrium to be reached in the neighborhood of the transformations. Heat contents of solid and liquid AuZn were measured between 488o and 1138°K by Kubaschewski.9 It was decided to determine the ordering energies of Au3Zn and to check on the possible disordering of AuZn by measuring the heats of formation of these alloys at several temperatures. The present paper reports measurements by liquid tin solution calorimetry of the heats of formation of the three phases of Au3Zn at temperatures within their regions of stability, and of AuZn at a low (322°K) and a high (800°K) temperature. EXPERIMENTAL Alloy Preparation and X-Ray Examination. Ten-gram ingots of Au3Zn and AuZn were prepared by melting weighed amounts of gold (99.95 pct Au) and zinc (99.99 pet Zn) together in sealed evacuated Vycor tubes at 1060°K, followed by rapid quenching. The resulting ingots were homogenized at 900°K for 10 days. The Au3Zn alloy lost no weight in preparation, the loss from the AuZn alloy was enough to introduce an uncertainty of only 0.1 at. pet in its composition. Filings for X-ray examination were taken from various parts of each ingot and strain-annealed at 723°K for 20 min in vacuo. The Au3Zn filings were quenched rapidly to retain the disordered phase. Diffraction patterns were determined with an X-ray dif-fractometer using copper Ka radiation.

Jan 1, 1968

By David W. Brunton

FROM the old-fashioned " grab-sample " to the modern timing- . device, which takes a machine-sample with mathematical precision, there is a wide gap which was only crossed' by many years of toil and unremitting endeavor. Even to-day, notwithstanding the advancement in the art, "grab-sampling" is still practiced-sometimes to afford the unscrupulous mine-promoter a basis for fairy-tales with which to entrap the too-gullible investor, and often by milling and smelting companies to determine the amount of moisture in custom-ores. The latter practice is almost as reprehensible as the former, and it causes more trouble and ill-feeling between seller and buyer than all other factors put together. No reputable concern to-day would think of attempting to determine by grab-sampling the amount of gold, silver, lead, or copper contained in an ore, and yet many buyers expect the miner to accept the results of grab-sampling in the determination of the amount of water contained in the ore, forgetting that accurate results are just as necessary here as in the determination of the metals, because the result determines the percentage of weight of the ore which shall be excluded and considered to have no value whatever. Samples for the determination of moisture should be taken with as great care as samples for the determination of. metallic content, and in order to avoid the extra expense of a separate operation moisture-samples should be taken from the sample-safe. As the sample reaches the sample-bin in a smaller stream and by a more circuitous route than the "reject" travels in its path to the outgoing car, it loses more moisture en route, and a constant should be added to compensate for this difference. Carefully-conducted experiments have shown that the difference in loss of moisture between the two routes does not exceed 10 per cent. in summer and 7 per cent.. in winter. For instance,

Aug 1, 1909

By D. Argyriades, G. Derge, G. M. Pound

The electrical conductance of molten FeS was studied as a function of temperature and composition. It was found that stoi-chiometric FeS (36.5 pct S) shows a minimum specific conductance of 400 ohm-1 cm-1. For the Fe-rich melts the conductivity increases with increasing amount of iron to about 4850 ohm-1' cm-I at 26 pct sulfur. For the sulfur-rich melts the conductivity increases with increasing amount of sulfur until it reaches a maximum of 1300 ohm1 cm-1 for 37.4 pct S and decreases to 800 ohm-1 cm-1 for 38.3 pct sulfur. It is concluded that molten stoichiometric FeS is an intrinsic semiconductor in which the forbidden gap in the electron energy level diagram is quite narrow. Excess sulfur acts as an acceptor impurity in providing a component of positive hole conduction, while excess iron acts as a donor impurity to increase the n-type conduction. K. Bornemann and G. von Ftauschenplatl used a four-terminal, direct-current cell to study the electrical conductance of molten copper sulfide. Their measurements showed a high conductivity, of the order of 100 ohm-1 cm-" and a positive temperature coefficient. From electrolysis studies, Savelsberg2 concluded that molten Cu2S and molten FeS are electronic conductors. Further. he found that molten FeS has a high specific conductance, of the order of 1500 ohm-' cm-', and a small negative temperature coefficient. Pound, Derge, and Osuch,3 using an ac Kelvin circuit and a four-terminal cell, measured the specific conductance of molten Cu,S, molten FeS, and various mixtures. They found that the specific conductance of the solutions followed a roughly additive rule of mixtures. Yang, Pound, and Derge4 showed from electrolysis measurements that the mechanism of electrical conduction in molten Cu2S, FeS, and mattes is primarily electronic. Further, they observed that small additions of CuCl suppress the electronic conduction whereupon the system behaves ionically.

Jan 1, 1960

Jan 1, 1971

By H. l. Burghoff, J. S. Porter

This paper presents the results of an investigation of the effect of cold-working and annealing treatments upon the occurrence of preferred orientation in annealed brass wire. The subject has received less consideration than has the same phenomenon in annealed brass sheet or strip, and perhaps very properly so in view of the obvious and undesired results obtained in the forming or drawing of strip having preferred orientation, which results are in contrast with the lack of any spectacular visual evidence in wire. The factors of cold-working and annealing that control preferredness in strip are naturally expected to function similarly in wire. Indeed, one might expect the condition to be commoner in annealed wire, because the severity of the cold-working between anneals in commercial Practice is usually much greater for wire than for strip. For example, a wire that is to be finished in the annealed condition on a diameter of 0.040 in. is commonly subjected to a final reduction of 90 per cent between the last two anneals, the large reduction being accomplished in the successive drafts of a multiple-die machine. Wire for rivets, for instance, on 0.125-in. diameter, although given a slight reduction of about Io per cent after the final anneal, may have been given a prior reduction of 75 to 80 per cent. Barrettl states the fiber axis of annealed 70-30 brass wire to be [111], with some indication of [IOO] as a secondary fiber axis. Croft and Sachs2 have reported briefly on some differences in brass wire with random and with preferred orientations. They showed statistical plots of the twin-band angles with respect to the wire axis or direction of drawing as observed on an axial plane. They also pointed out that the resistance of the wire to corrosion-cracking when externally stressed and exposed to a mercurous nitrate solution decreased as the degree of the preferred orientation increased. Material and PRocedure All the wire for the investigation was processed from one portion as cut for extrusion from a cast billet of the following analysis: 70.13 per cent copper; 0.01 lead; 0.01 iron; 0.02, nickel; 29.83 zinc (by difference). The billet section, 8 in, in diameter, was extruded hot to 3/4-in. diameter, after which it was processed by alternate cold-drawing and annealing treatments as indicated in Fig. I. The earliest anneals in process were at about 1100°F. and the next to the last reductions by drawing were 58 per cent. The diameter of the finally annealed wire was 0.132 in. Semifinal anneals were at 800°F., 950°F. and 1100°F. and reductions of 47.5, 74, 83.5 and 89.5 per cent were employed in drawing to 0.132 in. The final anneals were at 87s°F., 1050°F. and 1250°F. and were

Jan 1, 1944

By H. l. Burghoff, J. S. Porter

This paper presents the results of an investigation of the effect of cold-working and annealing treatments upon the occurrence of preferred orientation in annealed brass wire. The subject has received less consideration than has the same phenomenon in annealed brass sheet or strip, and perhaps very properly so in view of the obvious and undesired results obtained in the forming or drawing of strip having preferred orientation, which results are in contrast with the lack of any spectacular visual evidence in wire. The factors of cold-working and annealing that control preferredness in strip are naturally expected to function similarly in wire. Indeed, one might expect the condition to be commoner in annealed wire, because the severity of the cold-working between anneals in commercial Practice is usually much greater for wire than for strip. For example, a wire that is to be finished in the annealed condition on a diameter of 0.040 in. is commonly subjected to a final reduction of 90 per cent between the last two anneals, the large reduction being accomplished in the successive drafts of a multiple-die machine. Wire for rivets, for instance, on 0.125-in. diameter, although given a slight reduction of about Io per cent after the final anneal, may have been given a prior reduction of 75 to 80 per cent. Barrettl states the fiber axis of annealed 70-30 brass wire to be [111], with some indication of [IOO] as a secondary fiber axis. Croft and Sachs2 have reported briefly on some differences in brass wire with random and with preferred orientations. They showed statistical plots of the twin-band angles with respect to the wire axis or direction of drawing as observed on an axial plane. They also pointed out that the resistance of the wire to corrosion-cracking when externally stressed and exposed to a mercurous nitrate solution decreased as the degree of the preferred orientation increased. Material and PRocedure All the wire for the investigation was processed from one portion as cut for extrusion from a cast billet of the following analysis: 70.13 per cent copper; 0.01 lead; 0.01 iron; 0.02, nickel; 29.83 zinc (by difference). The billet section, 8 in, in diameter, was extruded hot to 3/4-in. diameter, after which it was processed by alternate cold-drawing and annealing treatments as indicated in Fig. I. The earliest anneals in process were at about 1100°F. and the next to the last reductions by drawing were 58 per cent. The diameter of the finally annealed wire was 0.132 in. Semifinal anneals were at 800°F., 950°F. and 1100°F. and reductions of 47.5, 74, 83.5 and 89.5 per cent were employed in drawing to 0.132 in. The final anneals were at 87s°F., 1050°F. and 1250°F. and were

Jan 1, 1944

By Horace R. Graham

CHEMICAL nitrogen and the "nitrates" of commercial significance are derived mainly from three basic sources: (1) the natural deposits in the form of nitrate-bearing earth and clay, which, being largely water-soluble, can exist only in the most arid portions of the earth; (2) coal, which yields not only nitrogen but also the additional by-product hydrogen required for the synthesis of ammonia and its compounds; (3) the atmospheric air, which contains 75.5 pct by weight of pure nitrogen gas. NATURAL NITRATE DEPOSITS Occurrence Although natural nitrate deposits consisting principally of the salts of sodium and potassium have been reported in Egypt, South Africa, Mexico, Argentina, Colombia, Peru and the USA, the only known natural nitrate resources of commercial importance are the vast deposits found in the Atacama, Tarapaca and Antofagasta regions of northern Chile, lying in a narrow strip 10 to 50 miles wide and roughly 450 miles long, between latitudes [19O] and [26OS]. In addition to yielding nitrates of sodium and potassium, these deposits are the source of the major portion of the world's supply of iodine. The nitrate-bearing ores known in Chile as caliche are variable in composition and may contain from 5 to 30 pct each of nitrate, chloride and sulphate. While the deposits are present chiefly in the form of the sodium salt, potassium, magnesium and calcium also occur in percent- ages from 0.1 to 5; usually as double salts with sulphate, which are but slightly soluble in nitrate leach solutions. These double salts are bloedite(Na2S04.MgS04.4H,0), glauberite (Na2S04.CaS04), polyhalite (K2SO4.- MgS04.2CaS04.2H20). Some deposits contain up to30 pct of their ni-

Jan 1, 1949

By Paul F. Kerr

The uranium-producing areas near Marysvale, Utah provide an unusual group of veins and replacement deposits associated with a Pliocene-Oligocene intrusive and extrusive igneous complex. Aside from several scattered occurrences in older volcanics, uranium-bearing veins occur in intrusive quartz monzonite or granite and in nearby rhyolite. The veins in general trend within a few degrees of N55-65°E, are more or less vertical, and are found along the western margins of a quartz monzonite intrusive. The replacement deposits are found in rhyolite intrusives and also in the lower portions of rhyolite flows close to places where veins forming solutions rising through quartz monzonite came into contact with the rhyolite. At least 14 veins have been mined since uranium mining operations began at Marysvale some 16 years ago, while ore has been extracted through about 10 miles of underground workings. The active mmmg area is approximately 3000 feet long and in the central portion 1500 feet wide. Ores have been found over a vertical spread of almost 2000 feet, between the highest occurrence on Jungfrau Hill ( ca. 7200 feet) and the deepest discovery of ore in a drill hole near the Prospector Mine (ca. 5200 feet). Under existing mining and marketing conditions, mining along the veins in the main area appears to have reached an economic limit at about 500-600 feet below the surface. However, orebodies in the rhyolite and in the upper portions of veins close to the rhyolite are subject to active mining. Isotopic dates indicate that the uraninitebearing veins and replacements are late Miocene or Pliocene (ca. 10-13 m.y.), the rhyolite is Miocene (avg. 17.5 m.y.), and the quartz monzonite and granite intrusives are late Oligocene (23.6 to 26.0 m.y.). The older Bullion Canyon Volcanic Series which encircles the area is Oligocene (28 to 31.1 m.y.). Two generations of alteration occur: ( 1) Earlier alunitic veins and replacement bodies are found in the Bullion Canyon Volcanic Series, none of which have produced uranium; (2) Later argillic alteration associated with the introduction of uranium is found along veins, replacing glassy rhyolite dikes and replacing phases of the rhyolite.

Jan 1, 1968

By Robert A. Kinzie

PAGE Introduction.............334 Climate.............335 History............335 Geology.............341 Mining..............343 Shafts............343 Stations and Ore-Bins....345 Levels. Drifts. Cross-Cuts, Winzes. etc........347 Stoping Surface-Pits,......... 351 Underground Stoping........... 332 Chutes. Drilling. etc.....354 Tramming............. 358 Sampling. Maps and Assaying.........360 Labor.............361 Machinery.............362 Milling,..............362 Crushers.............366 Battery-Foundations...........368 Power (Water—Coal)...........373 Methods of Catching the Free Gold...374 Screens..............376 Apron-Plates,.......376 Concentration..........378 Clean-Up,.............378 Labor..............382 Costs..............382 Introduction. The Treadwell Group comprises five mines. four of which are situated on the northeast side of Douglas island. directly opposite the town of Juneau, and one about four miles back of Juneau. These mines are owned by four separate companies. but are all under the same general management. Those situated on Douglas island are : The Treadwell mine. owned by the Alaska Treadwell Gold Mining Co.. the Mexican mine. owned by the Alaska Mexican Gold Mining Co.; and the Ready Bullion and Seven-Hundred-Foot mines. owned by the Alaska United Gold Mining Co. The Juneau mine. owned by the

Jan 1, 1904

By Frank Firmstone

IN the early decades of the past century the method of starting iron blast-furnaces by "scaffolding-down" seems to have been in universal use for coke-furnaces and, at least in this country, for charcoal-furnaces also. It was likewise used at some of the early anthracite-furnaces, as is stated in the important "Reminiscences" printed by the late Samuel Thomas. I It had been generally superseded in Great Britain as early as 1850, by some variations of the plan now used, and was rarely used when Percy wrote.' On the Continent it was still followed in the 60's,3 and was elaborately described by Valerius in 1852. I have never read any statement as to when and by whom '4 scaffolding" was first dispensed with, but the following extracts from journals of my father seem to fix an authentic date for an instance in this country:

Mar 1, 1907