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By AIME AIME

IT IS not necessary to travel hundreds of miles from New York City to see examples of modern mining methods applied in tunnel work. In the various subway jobs, for instance, underground work is done on so large a scale as to justify the oft-quoted remark that here is the world's greatest mining camp. In such cases, however, the size of the excavation makes wholly disproportionate comparison with the haulage drift or adit level of the West. But in West Orange, N. J., Antonio -DiMarco is driving a tunnel from two headings on opposite sides of First Mountain, under contract with the town of West Orange, which is similar in nearly every respect to the type of work that must be done at the mines, and merits description as an example of fast and economical tunnel driving under conditions commonly met with by the mining engineer in more inaccessible parts.

Jan 1, 1931

Roger A. Markle was sworn in as director of the US Bureau of Mines on September 27, 1978, ending a 20-month period during which the Bureau was without a leader. Some Bureau observers wondered out loud if this delay was a signal that Interior was slowly pulling the rug out from under the agency. When Markle was finally tapped for the director's position-coal industry executive to lead an agency that had been stripped of its fossil fuel R&D functions as well as its safety and health enforcement role--the speculation was fast and furious: One camp had it that Markle was brought in to phase out the Bureau; others said that Markle's directive was to revitalize the organization by charting a new Bureau course. In March, with the proposed Department of Natural Resources making headlines, ME went to Washington to interview Markle about his first six months as director and the goals he has set for the Bureau's future.

Jan 4, 1979

By Vernon Read

Among the technical problems in AEC's Project Plowshare is the necessity of determining the precise effects of nuclear blasts in all kinds of ground formations. This calls for extensive in-situ sampling with full core recovery. In the exploration drilling that often precedes a new test detonation, the AEC and its contractors have applied both novel and established sampling techniques in many new ways, giving improved core recovery and reduced costs in all subsurface sampling and testing. Unique among these new techniques is one that achieved full recovery and virtually undisturbed samples from near-surface layers of unconsolidated boulder-strewn alluvium. This approach involves a combination of oilfield and mining core drilling equipment. Oilfield heavy drill collars, rotary stabilizers and thick, vibration-reducing oil emulsion drilling mud were used in combination with the precise hydraulic feed control, fast (over 300 rpm) bit rotation, light drilling weight, hard rock core barrels and multi-step diamond bits that are common in mining industry diamond drilling.

Jan 8, 1963

By R. W. Raymond

(Cleveland Meeting, October, 1912.) UNDER the names of Conservation, Social Justice, the New Nationalism, and Progressive Democracy, many earnest reformers are calling for a new system of Federal government to replace the one which they ascribe to our fathers, and declare to have been outgrown. They are proposing the lease by the United States of certain natural resources on the public domain, for the profit of the U. S. Treasury, instead of the sale of such resources to private citizens or associations of citizens. In the present paper, I shall try to show that this notion is neither new nor good. It is clear that the United States is owner, as well as sovereign, of public lands acquired through conquest or purchase from other nations, as it was owner of the public lands expressly ceded to it by the original thirteen States. Its policy with regard to such lands involves, therefore, two entirely distinct questions: What shall it do as owner ? and, What shall it do as sovereign ? The latter question, it must be confessed, has played but a small part in our history. All the prerogatives of sovereignty which were not explicitly conferred upon the Federal government by the original States were retained by them. The ownership or control of precious or other metals or minerals in the earth, for instance, was thus reserved to the several sovereign States. New York, South Carolina, and other States have passed laws on this subject, without regard to Federal authority. As new States have been successively admitted to the Union, the tacitly accepted theory has been,. that they thereby acquired all the reserved rights of the original States. These newer States largely outnumber the first. thirteen; and it is inconceivable that they would submit to any deprivation of sovereignty based on historical grounds. One of them-Texas-came into the Union by its own voluntary

Oct 1, 1912

By J. W. Smith

Our continuous search for underground productivity improvements has been brought about by the diminishing ore grades in existing underground mines. The need for more efficient mining methods is a result of the economic problems facing our industry today, and this has caused us to evaluate underground haulage methods which have traditionally been the "bottleneck" in the flow of material from the ore in the natural state to the surface processing facility of any underground mining operation. Small improvements in the face haulage systems have yielded much greater benefits as they relate to overall mine productivity so it's only natural that we are all concerned with the best method of moving ore from the face to the main line haulage. In a recent paper titled "Underground Haulage Trucks - Gaining Momentum Worldwide", Richard A. Thomas concludes that the use of trucks to haul ores in underground mines is on the increase spurred by the convergence of a number of technology advances and economic realities. Perhaps the most important stimulus for the growth of trackless haulage is the high degree of haulage flexibility in underground operations. On the economic side, the demand for higher productivity from underground mines has resulted in larger physical dimensions of haulage roads, that is, higher backs and wider drifts to provide more room for high capacity haulage units. In the process of determining the most effective type of equipment for haulage, the power source must be a major consideration. For the purpose of this paper, we will limit the comparison to rubber-tired trackless haulage vehicles and not try to make a comparison between rubber-tired haulage, continuous haulage systems and rail-mounted haulage. Cost is perhaps the only really measurable factor when making a comparison between electric and diesel haulage. You will find that some costs will be very well defined in absolute terms. In other areas of comparison, cost can be fairly well estimated, and yet in still others, the costs are totally arbitrary. Let's take a look at some of the cost considerations. (Figure 1) first of all, is the initial cost of the equipment. This capital cost quite often is a determining factor in the type of haulage vehicle to be selected, yet this initial cost is perhaps the most insignificant of all costs when evaluating an operation over the long term. Of much greater concern, is the cost of maintenance. This cost will often run three times the original capital investment during the life of a single piece of haulage equipment. This factor can include rebuild to extend the life of the original capital investment, but certainly includes the labor and materials necessary, plus the inventory to keep the equipment in good repair. Perhaps one cost which is now playing an even greater role in the rubber-tired haulage operation, is the cost of fuel. Conoco has recently come up with some rough estimates which indicate that diesel fuel will cost an average of three times the equivalent kilowatt output in direct electric power. Diesel fuel is almost twice the cost of stored electric power. (This of course relates to the efficiencies of charging and recovery of power from lead acid storage cells.) These particular figures of course will vary from one area to another but I think that there is enough significance here to certainly warrant the further study of fuel costs for each particular area or mine. Another cost is breakdown expense. This must be treated differently from maintenance costs because a potentially larger expense is involved, more than just parts and labor. Now we have to deal with the cost of lost production time, which can have a much greater overall effect. Mine plan economics are another cost consideration where we can't make a comparison without looking at specifics. Here you must look at the movement of power centers vs. the flexibility and freedom of movement of vehicles. The determination must be made as to what types of equipment will fit into any predetermined mine plan and if a change in the planned roadway dimensions for the mine plan itself would be more economical so that more efficient type of equipment could be utilized. Finally, two of the most important aspects to be considered with potential ramifications far beyond what we have mentioned previously, is the cost of health and safety, which is really the cost of meeting current and future government regulations, reasonable or otherwise. And of course, when making any consideration here it is impossible to come up with anything more than an educated guess on the cost of meeting the new regulations. Now let's take a look at some of the advantages of diesel vehicles as well as advantages offered by electric vehicles, both battery and cable powered versions (Figure 2). Much of the data used in this comparison is based on experience with three vehicles manufactured by Jeffrey Mining Machinery Division, Dresser Industries. Jeffrey manufactures all three types, each with approximately a 15-ton capacity, even though few of these Jeffrey vehicles are used in uranium mining operations. Much of our experience comes from the 4114 diesel powered RAMCAR which is a 4-wheel drive, articulated steering,vehicle powered by a Caterpillar 3306NA engine and using a powershift transmission. This will be compared with the performance of the Jeffrey 404H battery powered RAMCAR with articulated steering which utilizes a separate 35 HP DC drive motor on each of two wheels with solid-state speed controls, and the final comparison will be made on the Jeffrey 4015 cable-reel shuttle car which is powered by two 60 HP constant

Jan 1, 1982

By Scott Turner

AS IN OTHER lines of engineering, progress in mining was influenced during 1931 by the world-wide economic depression. Low-metal prices ? resulted in active efforts to reduce production costs of base-metal mines. Probably the chief factors in reducing costs were the greater efficiency of labor, due to better relations between management and workmen, to the scarcity of jobs, reduced wages in most base-metal mines, and to lower overhead expenses secured by drastic economies. The use of mechanical equipment continued to increase, due to more efficient designs and a better practical understanding of the applicability of such equipment to requirements. With one exception, there were no startling innovations in stoping methods, but there was notable progress in the better adaptation of mining methods. In the United States, no large new base-metal or silver mines came into production during the year. At a number of mines, output was curtailed, and operations were discontinued at several important producers. On the other hand world-wide interest in gold was stimulated, as evidenced by increased output from producing mines, the reopening of some old properties, and intensive prospecting for new deposits. Mining methods at several gold mines in the United States and Canada were modified or changed to obtain a higher extraction of ore or to meet the need for additional support of the mine workings at depth. Permanent record of improvements in mining practices has been made during the year by the Federal Government; the industry thus has available for its use more detailed information than ever before. This work was done through cooperation of the mining companies and their representatives with the United States Bureau of Mines, whereby many papers descriptive of practices and costs at individual mines have been published. The Bureau's program of describing methods and costs at the more important mines (those producing in excess of $100,000 annually) is nearly completed, but papers will be issued concerning such additional mines as may possess particular interest because of novel methods, exceptional conditions or operating results, and others to supplement reports already issued. In this way, an interesting record of the results of changing conditions, or innovations in methods, will be made available.

Jan 1, 1932

By AIME

Rapid exhaustion of iron ore in Minnesota's Mesabi range has led the Jones & Laughlin Steel Corp. to develop the world's largest open quarry mine for magnetite iron ore, at Star Lake, N. Y. Already contributing more than one eighth of the total iron ore used by the Corporation, the expanding Benson mines are destined to become Jones & Laughlin's important source of ore within the next ten years. In 1947 over 650,000 long tons of high-grade ore was produced. Output this year is expected to reach 800,000 long tons and studies are under way to increase production further. The future of the Benson mines is assured as present indications show large reserves of ore. The property, briefly described in M & M for November 1943, then under

Jan 1, 1948

By P. L. Weston, S. E. Khalafalla

A systematic study was made by the Bureau of Mines on the effect of so me hypothesized accelerators for the process of wustite reduction in carbon monoxide. When small concentrations of promoter materials in the order of 0.69 at. pct were added to the reducible charge, the rate of reduction to iron was increased. Promotion phenomenon prediction was made in light of a suyface reduction mechanism with the aid of Vol'kenshtein's effect regarding the propagation of crystal lattice disturbances by small amounts of relatively larger interstitial ions. The acceleration produced by a typical promotor, such as potassiunl, increases with protnoter concentration up to a maximum, beyond which the reduction rate decreases. Concentration for maximum promotion depends on the nature and physicochemical properties of the promoter. The extent of reduction rate enhancement is found to be directly proportional to the atomic volume and electronic charge of the additive. DESPITE the enormous volume of literature on iron oxide reduction, very little is reported concerning additive or impurity effects on this important metallurgical process. The beneficial effect bf calcium compound additions on the reducibility of iron oxide sinters has been reported by Tigerschiold,1 vor dem Esche,2 and Edstrom. Doi and Kasai~ found that the addition of lime or limestone to iron ores helps to break up any unreducible compounds, such as fayalite or ilmen-ite. and thus free the combined iron for reduction. Schenck et al. 5 suggested that the increased reduction rate obtained when adding lime could be accounted for by the instability of wustite in the presence of lime. Acid-base slagging reactions resulted in wustite disproportionation according to The dicalcium ferrite formed will yield iron and calcium oxide during reduction. Regenerated calcium oxide dissociates more wustite. This mechanism has been used by Seths and white7 to explain their experimental results. Recently, Strangway and ROSS' attributed the calcium carbonate acceleration of iron oxide agglomerate reduction to increased porosity, both initial as well as that developed during reduction. Aside from calcium carbonate, or oxide, no other promoter was noted in the literature, except for a brief mention by Barrett and woodg on the effect of sodium carbonate and aluminate as activators for the hydrogen reduction of magnetite at 600°C. The present investigation systematically studied a host of other promoters, including calcium and sodium, in an attempt to elucidate the mechanism by which promotion takes place and to fit the results into a simple chemical model. To attain this goal, the effect of promoter physical properties, such as atomic volume, electronic charge, and concentration are related to wustite reduction kinetics in this paper. Wustite reduction to iron, rather than the overall hematite reduction, was chosen since this reaction is known to be the slowest, and hence the rate-deter mining step for the overall iron oxide reduction process. EXPERIMENTAL PROCEDURE Raw Materials and Their Preparation. The pure or impregnated wustite pellets were prepared from minus 400-mesh chemically pure hematite powders. A known weight of hematite was thoroughly and uniformly mixed with a calculated weight of the additive. The mixed paste containing 35 wt pct water was gradually heated from 400" to 1200° C and fired at 1200°C for approximately 4 hr in an air atmosphere. After cooling, the sinter was pulverized to minus 100 mesh and pelletized into minus 4- plus 5-mesh spheres. Pellets were fired, similarly to the paste mix, air-cooled, sized, and stored. An appropriate weight of the charge (20 g) was placed in a zirconia reduction tube maintaining a uniform oxide bed height of 1 cm and a cross section of 7.1 sq cm for all of the test runs. The samples were supported in the vertical reaction tube by a bed of fragmented insulating firebrick plus 3- to 6-mesh alumina beads. The hematite was then transformed to wustite by reduction with a 30 pct CO2-70 pct CO gas mixture at 1000°C in a globar furnace. Complete conversion to wustite was ascertained by a continuous infrared gas analyzer recording the CO-CO2 content of the effluent gas until no carbon monoxide was absorbed from the inlet gas, and inlet-outlet gas analysis remained constant for 30 min. The wustite sample was then reduced with 100 pct CO at 100WC. From the recorded data, an initial rate of reduction was determined by the initial slope of the graph percent reduction vs time. In order to estimate the accuracy of the data, five separate determinations of the reduction curve of pure wustite, under otherwise identical conditions, were performed. The maximum deviation from the average reduction at 14 min amounted to 2 2 pct reduction. This deviation corresponds to 3.8 pct variation based on the percent reduction of the sample. Aiiy impurity effect below the limits of this maximum deviation was considered a spurious result. If the effect exceeded + 4 pct, then it was considered as a positive one. Considerable care was exercised in determining the initial rate from the slope of the initial segments of the curve. Each reduction curve was examined separately on large graph paper and the best tangent to the curve at zero time was drawn. The slope of this tangent was taken as a measure of the initial fractional reduction per minute. Although the time required to reach 50 pct reduction may be of special practical significance, initial rate measurements are invaluable in fundamental studies. These rates provide a measure of the process kinetics on the initial

Jan 1, 1968

By AIME

Cores must be cut in half to secure workable samples for saturation, porosity, and permeability tests. They come from the field labeled as to the type of core, well, and depth from which they came.

Jan 1, 1948

By V. C. Allen

THE Tri-State Zinc, Inc., Galena, Ill., was confronted with the problem of securing ore from a deposit because the hoisting shaft was several thousand feet from the mill. The orebody is several thousand feet long, averaging 200 ft in width and 60 ft in height and opened up by vertical shafts some 300 ft deep. Mining is by the room-and-pillar method. During the initial operation the ore was loaded by conventional electric 1/2-yd boom-and-dipper shovels and hauled to the shaft by 8-ton diesel trucks. This underground ore loading and hauling was well adapted to the conditions and productive of low costs per ton. However, with the mill situated as mentioned, a triple handling of all broken rock was necessary: l—from the stope to the shaft by truck, 2—up the shaft by skip br can into the surface hopper, and 3—by truck from the surface hopper to the crushing plant at the mill. In addition to the repeated handling, serious troubles were encountered during the winter because of freezing in the shaft hopper. Consideration was given to either moving the mill to the new orebody or to the construction of a second mill. The presence of other orebodies to be mined at a later date made the first alternative impractical while the capital outlay for a second mill, when the present plant of approximately 850 tons per day was deemed sufficiently large for the total reserves, made the second alternative also unwise. It was decided to retain the mill in the originals location and continue to move the ore to it. The idea of driving an inclined adit from the surface to the bottom of the orebody suitable for truck haulage and big enough to allow the passage of all mechanical equipment was conceived. Among the apparent advantages of such an incline were: 1— Direct haulage from the stope to the mill without rehandling. 2—Elimination of virtually all grizzlies. Trucking from underground to the mill would do away with all hoppers, chutes, gates, and skips and make the maximum rock size dependent solely on the size of the shovel dipper at the mine and the crusher opening at the mill. 3—Less secondary blasting would be needed. 4—Ease of transporting equipment and supplies underground. Shovels and trucks could be taken through the incline intact. 5—Equipment could be brought to the surface for repairs and servicing without loss of time. The same advantages of ease in moving would be present in the handling of men, steel, powder, and supplies. 6—There would be far less difficulty in increasing the amount of tonnage that could be moved by truck up an incline than would be found in attempting to increase the capacity of a shaft. 7—All the broken ore in the stopes would serve as bin capacity, as it would take the breakdown of all of the loading and hauling equipment to have the same effect as a delay in shaft hoisting. 8—All danger of men being trapped in the mine as a result of shaft fire or power stoppage would be eliminated. 9— Virtually all trouble from severe winter conditions would be eliminated by the direct haul underground to the mill. The decision was made to proceed with the driving of an inclined adit. The topography of the surface between the orebody and the mill was such that it was possible to locate the portal at a point 170 ft above the mine floor and 1800 ft horizontally from the central point of the orebody to the south and 2500 ft from the mill to the north. A grade of 10 pct was found to be optimum for continuous truck haulage when the various factors of speed, safety, and truck maintenance were all considered. The incline as driven was consequently 1700 ft long on 10 pct grade and 12 ft high by 17 ft wide in cross section. The tunnel-driving equipment was chosen so that it could be used in mining after the completion of the tunnel. Drilling was done with a jumbo with two Joy jibs mounting 3-in. drills, loading with an Allis-Chalmers diesel-powered, front-end loader of approximately 11/4-yd capacity, and hauling by Koehring Dumptor trucks of 8-ton capacity, diesel-powered. The width of the tunnel allowed the end loader and Dumptor to be placed abreast. Since the Dumptors can be driven either forward or backward with equal facility, loading was accomplished without turning around either machine throughout the loading operation. The crew in addition to the tunnel foreman was comprised of three men per shift at the start and in the later work, four men. Each crew could perform any part of the working cycle. If the drilling was completed and the round blasted in the middle of a shift, the same men would proceed with the loading and hauling. Since the mine already had been drained to the bottom levels, no water was encountered. At the halfway point the tunnel was widened for approximately 100 ft to permit trucks to pass. The total cost of the tunnel excluding the capital outlay for equipment, which was all continued in use in the subsequent mine operation, was $60,363.00 or $35.50 per ft. The tunnel was completed at the end of June, 1949 and has been in continuous use since that time. In the five months from July to November inclusive, 106,049 tons have been transported to the mill or an average of 835 tons per day. No unforeseen disadvantages have been encountered and the advantages which had been predicated before the adit's construction have been more than realized. As previously mentioned, the deposit is worked by the room-and-pillar system with occasional faces up to 125 ft high. Except in driving development drifts when diesel-powered, front-end loaders such as were used in the tunnel are employed, all shoveling is done by Yz-yd boom-dipper type shovels electrically driven. These units need a width of 25 ft and a height of 14 ft in which to operate. All hauling is by diesel trucks, mainly Koehring Dump-tors. Roads are maintained with caterpillar tractors and a road grader. The tonnage output from the

Jan 1, 1952

By S. Arthur Krom

(Wilkes-Barre meeting, June, 1911.) RECENT experiments indicate that the usual type of concentrating-table is not only poorly adapted to produce the desired results, but also is based upon an incorrect principle, namely, the use of riffles to perform the work of stratifying the various minerals. We have heard a good deal about riffles for concentrating-tables; exhaustive experiments have been made to discover the proper form of riffle, or to prove the superiority this or that form; disputes and patent-litigation have arisen over the matter of riffles, and thus many have been led to believe that the riffle is the saving-device upon which the process of concentration depends. In the present paper it is proposed to show that the riffle is greatly overrated as to the part it performs in the concentration of minerals, and that, in the near future, it may possibly be eliminated entirely from that process. The experiments in question have shown that the troublesome riffle can be considered of secondary importance at the most, and should be so classed in the construction of a concentrating-table built upon the right lines. The action of any form of riffle on a concentrating-table is such as to upset and retard the process of settling and stratifying the various minerals on the table-deck. From the deck of a riffled table no concentrates can be delivered until the deck is "bedded," that is, until a sufficient amount of metallic values have been fed to the table to spread over a large portion of the deck, forming a substratum of the heavier minerals. This substratum must be maintained by the feed within quite narrow limits, and directly proportional to the rate of discharge from the table, in order that the bed shall not be lost. On the other hand, the riffles having a very lim-

Jul 1, 1911

By John B. Hastings

THE reconnaissance of the Goldfield mining-district, described in this paper, was made in May and June, 1905, and, though this time was too short for a complete report, the work accomplished may serve as a basis for more thorough future research. Dr. Berkey has examined the rocks collected by myself, and I am using his classifications. As a result of further field-work, and Dr. Berkey's examination of the rocks, this paper is an elaboration and correction of my views expressed in the Gold-field Sun, May 12, 1905. The sketch-map, Fig. 1, illustrates the geology of the Goldfield mining-district. Goldfield is an eruptive complex, consisting of alaskite (binary granite), hornblende-andesite, hornblende-dacite, rhyolite, pyroxene-andesite, pyroxene-dacite (sometimes containing, a small quantity of olivine), quartz-felsite, olivine-pyroxene-ande¬site, and basalt. Considering this list to be arranged in the order of eruption, the age-sequence of the rocks conforms to Richthofen's law, formulated in 1868. From his observations in Europe and America, propylite, andesite, trachyte, rhyolite and basalt, occurring together, succeed one another. In a general way, the medium and more acidic rocks precede the most asic-a result supposed by some to have arisen from their relative position beneath the surface, and due to a rude stratification by gravity in the original cooling magma during the older history of the earth, or to the same differentiation occurring during a quiescent (molten ?) period preceding eruption, being aided by the ease with which acidic rocks, mixed with water vapors, become fusible. In effect, the more basic and less re-

Mar 1, 1906

The United States has never been a large coal exporting nation, and normally such shipments, aside from those to Canada, have been relatively small. Since the first railroad was built across the Allegheny Mountains, the imports have been quite small also, although fairly large tonnages of Canadian coal have moved into the extreme northeastern and northwestern states at times. In Table 21 are shown all data available about these shipments from 1790 to 1939, inclusive.

Jan 1, 1942

By A. Lubinski, C. R. Stewart, K. A. Blenkarn

Storage porosity has been considered one of the important pore geometry characteristics of heterogeneous-porosity limestones. Storage pores are only containers for fluids, in contrast to flow channel pores which both contain fluids and provide continuity for fluid flow. The concept of another geometry characteristic, "poro-striction", is presented as a second pertinent variable in describing limestone pore space. In simple terms, poro-striction is a measure of the flow resistance between storage and flow channel pores. Alternating-flow core-testing theory provides flow relationships which can be used to divide the pore space of heterogeneous-porosity media into flow channel and storage pores and to measure the "porostriction" of the latter. Experimental application of this theory to naturally occurring heterogeneous limestones shows that "porostriction" and the ratio of storage to flow channel pores can be estimated. Porostriction and porosity ratio are microscopic characteristics which should influence oil-recovery efficiencies during certain types of displacement processes. A knowledge of pore geometry should be valuable in designing or selecting the most effective oil-recovery process for heterogeneous limestones containing storage porosity. INTRODUCTION The primary and secondary recovery of oil from limestone (carbonate) reservoirs has been and will continue to be a major source of the world's supply of petroleum. Oil recovery from this type of formation is strongly influenced by the size, shape and arrangement of the pore spaces which hold the formation fluids. Therefore, it is expected that better knowledge of pore geometry would lead to design of more effective recovery processes. The concept of storage and flow channel pores has been presented as an important characteristic of limestone porosity.1 In this concept, storage pores are only containers for fluids, whereas flow channel pores are those pores which both contain and transmit the fluids. The work reported herein involves the concept of an additional pore characteristic called "porostriction". This characteristic can be visualized as the flow resistance between the storage and flow channel pores. This resistance is an important pore geometry characteristic because it determines how easily gas and oil can replenish fluids removed from the flow channel pores. Fatt and co-workers at the U. of California have studied the effect of storage porosity and a factor similar to porostriction on transient-flow response of models of porous media during a pressure build-up or drawdown test.55 These models consisted of porous sandstone cores with artificially simulated storage pores. It is the purpose of this paper to describe tests which permit the storage porosity and porostriction of actual reservoir core samples to be estimated. The tests used for this purpose involved not transient tests but, rather, an alternating-flow technique in which pressure and flow are varied according to a sine-wave relationship. This technique may offer advantages over transient methods. THE POROSTRICTION CONCEPT Before discussing alternating-flow (A.F.) tests in core samples, it is necessary to define the porostriction parameter. For this purpose attention is focused upon an elemental volume v representative of a porous medium having both flow channel pores and restricted-access storage pores. For such an elemental volume, it is considered that the storage porosity is combined to form a single pore and that the access into the storage porosity may be represented by a single capillary tube connecting the single pore to the flow channels of the elemental volume. Let the radius of the capillary tube be designated as Rc and its length as L,. When fluid flows into or out of the pore, the following equation (Poiseuille's law) relates flow rate q to pressure drop p, and the physical characteristics of the tube and the fluid of viscosity jx.

By Michael J. Fiannnaca, Michael M. Easdon

Disseminated epithermal gold metallization occurs within a breccia unit interpreted to be a submarine debris flow. Gold is associated with oxidized pyrite, fluorite, and jasperoidal silica. A minimum of 8 Mt (9 million st) of open-pittable mining reserves will be cyanide heap leached as run-of-mine ore at a rate of 0.9 Mt (1.0 million stpy 1. Careful evaluation of the combined results of successive sampling projects has established the pit plan, minable reserves, and process metallurgy of the operation. The sampling procedure utilized for a given test depended upon geological interpretation as well as the procedures, successes, and failures of previous tests. This paper summarizes the sampling and testing procedures of the development effort in light of the project goals and attendant problems in achieving them.

Jan 1, 1985

By George W. Small

The ore, as it is raised from the mine, has all average assay-value of $11 per ton, chiefly in the form of free gold. All the ore goes directly to the stamp-mills, of which there are two. The older and larger mill contains sixteen batteries of five stamps each, with one Frue vanner to each battery. The new mill has eight batteries of five stamps and two Frues to each battery. The large mill is driven by Leffel turbine wheels, with a pressure of eighty feet, and a consumption of 600 miners' inches of water. The smaller mill is driven by "hurdy-gurdy" wheels with a pressure of about 550 feet and a consumption of 150 inches of water. At both mills the tailings from the stamps pass over about 20 feet of plates on their way to the Frues. In each set of plates the first or upper one is copper, the rest are so-called silver plates. The bullion from the stamps is about 800 fine in gold and 200 in silver. The concentrates from the Frues average from 11/4 to 1 ; per cent. of the ore stamped. They very rarely exceed 2 per cent. I was unable to get the exact assay-value of the concentrates, but it is said to vary between $100 and $200 per ton. The concentrates are treated at the chlorination works at the rate of one hundred tons per month. The capacity of the works is somewhat greater than this, but as the supply of concentrates is limited it is not deemed advisable to work them up any faster. Care is taken to keep the concentrates always damp until they are put into the roasting furnace. If this is not done, a decomposition of the pyrites begins, forming lumps which do not roast, and which consequently cause a loss of gold in the residues from leaching. A Fortschaufelungsofen is used for roasting. Its dimensions, incloding fire-box, are 12' X 80'. The hearth is one continuous plane, but the charges, of which there are three in the furnace at one time, are kept entirely separate. The furnace-men call the three compartments, the "drying," the " burning," and the " cooking" compartments. In the middle, or "burning" compartment, the ore is spread

Jan 1, 1887

By James P. Kimball

THE object of the present memoir is to put on record some practical experiments by the writer in smelting a silicious native magnetite with no other flux than the silicates of its own gangue: This work was undertaken in the spring of 1878 for Messrs. Naylor & Co., of Boston, to whom I am indebted for the privilege of here communicating its results, these having hitherto only been simply announced.* So far as I am aware these are the first instances of the kind in this country that non-calcareous; or rather non-carbonated, magnetites have successfully been used strictly alone as self-fluxing ores in the blast furnace; that is, without the addition of calcareous flux in the form either of limestone, quicklime, calcareous ore, or basic (calcareous). slag. Previous to the spring of 1878, when the Chateaugay magnetic ore was first brought to my notice, its use had been limited to Catalan forges, after "separation" at the ore bed. The blooms and billets made from it in the several forges of Messrs. Weed & Williams, and of Mr. Williams, were well known in the trade as among the best in the market. The Chateaugay ore beds proving of great magnitude, and the ore exceptionally low in phosphoric acid, it commended itself for blast-furnace purposes to Messrs. Naylor & Co., who employed nee to blow-in the charcoal furnace at Pittsford, Vt., on this ore in the spring of 1878., a previous trial of this ore in the same furnace in the fall of 1877 having proved abortive. This was therefore the first time that Chateaugay ore was successfully smelted in a blast furnace, although since well known as a blast-furnace ore possessing superior and exceptional qualities. The Pittsford furnace, although rebuilt in 1844, and subsequently improved by the former owner, the late J. Prichard, is one of the few furnaces of the old type with open breast (Sumpfofen) still active in this country. Its improvements on furnaces of the period, * Bulletin American Iron and Steel Association, XII, 1878, 132, 156.

Jan 1, 1881

By M. J. Marcinkowski, Gordon E. Lakso

An extensive investigation has been made of the deformation behavior associated with the Fe3Si super-lattice using transmission electron microscopy techniques. Above 243°K the stress-strain curve exhibits three stages. Stage I occurs at a very low stress level and is related to the generation of perfect superlat-tice dislocations. Stage II is characterized by an extremely rapid rate of work hardening and is associated with the Taylor type locking of these superlattice dislocations. Finally Stage III is related to dynamic recovery processes since the work hardening rate is very small. Below 243ºK, only Stage I is observed, but it occurs at a much higher stress level. This latter observation is related to the generation of imperfect dislocations in Stage I with the consequent production of second nearest neighbor antiphase boundaries. The reason for this is that insufficient thermal energy is available at these low temperatures to generate the complete and perfect superlattice dislocations. It has been shown that the fully ordered FeCo alloys, i.e., those possessing the B2 type structure, exhibit three distinct stages of work hardening whereas the corresponding disordered alloys show only one.&apos;" This difference in behavior between the disordered and ordered alloys has been attributed to the fact that dislocations in the former case travel only as ordinary 1/2ao(111) types whereas in the latter case the move through the lattice as coupled 1/2a0(111) dislocations separated by an antiphase boundary (APB), i.e., the so-called superlattice dislocation. Although some preliminary work has been carried out concerning plastic deformation in ordered alloys possessing the DO3 type superlattice,3 no detailed analysis similar to that described in Refs. 1 and 2 has been attempted. Specifically, it has been suggested that the superlattice dislocation in this particular type structure should consist of four ordinary 1/2ao<111> types bound together by first and second nearest-neighbor APB&apos;s. Fe3A1 and Fe3Si are the two classic alloys possessing the DO3 type lattice; however, because of the somewhat higher ordering energies associated with the FesSi alloy, which in turn assures that dislocations will travel through the lattice as perfect superlattice dislocations under at least some conditions, it was chosen for the present investigation. Because of the extreme brittleness of Fe3Si, all deformation was done in compression. Stress-strain curves were obtained using both polycrystalline samples as well as single crystals. In the latter case the crystals were oriented so that deformation could be controlled either by single or double slip. They were then wafered parallel to and at various angles to the operative slip planes. These wafers were in turn examined by transmission electron microscopy (TEM) techniques in order to determine the extent of the interaction from the dislocation configuration contained therein. EXPERIMENTAL PROCEDURE The alloys used in this investigation were arc melted under helium from electrolytic iron of greater than 99.90 wt pct purity and transistor grade silicon of 99.99 wt pct purity. A typical analysis of interstitial impurities showed 120 ppm 0, 15 ppm N, and 65 ppm C Because of the extremely low ductility of the Fe3Si alloys, it was necessary to spark cut 0.230-in. diam polycrystalline cylinders 0.400 in. long from arc-melted fingers using a thin-walled brass tube as a cutting tool. The polycrystalline alloys could not be recrystallized since very little strain was induced in preparation. However they were annealed at 1273°C for 15 min in evacuated vycor capsules to relieve any cooling stresses that may have developed during solidification and then air cooled. The resulting grain size of the alloy was 0.50 mm. According to warlimont4 1273ºC is just within the single phase field where FesSi possesses the DO3 type lattice. In addition because of this high critical ordering tem-ature, air cooling from this temperature was believed sufficient to fully order all of the Fe3Si samples used in the present investigation. For the same reason, no attempt was made to achieve any degree of disorder by quenching. In fact, rapid quenching from 1123°K caused cracking. Such cracking was first suggested by sato5 with respect to the experimental observations of Glaser and Ivanick.6 Single crystal compression specimens were spark cut from single crystal ingots grown in a Bridgman type furnace. The iron and silicon for the crystals was prealloyed by arc-melting two 130-g buttons which were cut into small pieces before remelting in the furnace. This procedure resulted in a long-range inhomogeneity of 0.5 at. pct Si between the top and bottom of the 2-in.-long single crystal ingot, which was assumed to be negligible in the present investigation. The single crystals, after orienting and spark-cutting, were about 0.37 in. by 0.37 in. in cross section and about 0.5 in. long. True stress-strain curves were obtained using an Instron Tensile Testing machine in conjunction with techniques described previously. 1,7 The strain rate was 0.05 in. per in. per min. Prior to testing, the ends of all the compression cylinders were hand polished using a special jig to insure parallelism after which the sides of the samples were electrochemically polished to eliminate stress risers and to facilitate slip line observations. Test temperatures between 77" and 823°K were obtained using various cooling and heating media as described in Ref. 7 while at the upper end of this temperature range, a mixture of equal

Jan 1, 1970

By K. H. Heino

Two 200 T/da Lurgi Turbulent Layer fluid-bed roasters have been in operation at Canadian Electrolytic Zinc Limited since 1966. Zinc concentrate containing 52-54% Zn, 9-11% Fe,31-33% S is treated to produce calcine containing 60-62% Zn,10-12% Fe, .l-.3% S as sulphide and .8-1.2% S as sulphate. Each roaster train comprises a waste heat boiler, two cyclones in parallel and a two-field electrostatic precipitator. Distribution of calcine collected is 10-15 percent from bed overflow, 55-65 percent from boilers, 20-25 percent from cyclones and 2-6 percent from precipitators. A Monsanto designed contact plant produces acid at a rated capacity of 365 T/da (100% H2304 basis) from roaster gases diluted to 7.0-7.5% S02. The process steps are gas cleaning and cooling by weak acid scrubbing, acid mist removal in electrostatic precipitators, gas drying in 93.5-96.5% H2304, conversion of sulphur dioxide to sulphur trioxide in a four-pass converter containing vanadium pentoxide catalyst, and absorption of the sulphur trioxide in 98.1-98.3% H2S04. Product acid is 93.2% H2S04. Conversion efficiency is 97-98 percent and overall sulphur recovery from concentrate is 87-90 percent. The adjacent fertilizer plant has a capacity of 135 T/da of P205. The products are 54 percent phosphoric acid, run-of-pile triple superphosphate, and diammonium phosphate. The plant is of Ugine-Kuhlman design and consists of an air-cooled multiple tank reactor, a tilting pan filter, and a forced circultation vacuum evaporator. Triple superphosphate is made on a continuous belt and partially cured in a rotary dryer. Diammonium phosphate is produced by ammoniating phosphoric acid in a single stage reactor. The slurry of ammonium phosphate is granulated in a blunger and dried.

Jan 1, 1970

By R. F. Domagala, M. Hansen, D. J. McPherson

Iodide zirconium was combined with calculated amounts of nitrided zirconium sponge and arc melted to prepare alloys in the 0 to 6 wt pct N region. Annealing treatments were carried out at 21 temperature levels. Metallographic examination of the heat-treated specimens permitted construction of the binary phase diagram from 0 to 6 pct N. Features of the diagram include the peritectic formation of both a and ß solid solutions. The maximum solubility of nitrogen is 0.8 pct in 8 zirconium and 4.8 pct in a zirconium. An X-ray study of nitrided materials was made in the range 6 to 13 wt pct N region because serious nitrogen losses were experienced when attempts were made to arc melt these high nitrogen alloys. PHASE relationships in the Zr-N system have been determined. Due to the inability to retain nitrogen in nitrogen-rich alloys subjected to arc melting, definitive work was possible only in the range 0 to 6 wt pct (0 to 30 atomic pct) N. Sufficient complementary X-ray work was done to permit construction of the binary phase diagram up to 13.3 wt pct (50 atomic pct) N (ZrN). Small ingots were prepared by arc melting master alloys with enough zirconium to produce an alloy of the desired composition. Following pretreatment, alloys were annealed at temperature levels between 600" and 2020°C. Determination of the phase boundaries was then accomplished by metallographic evaluation of specimens quenched from the various temperatures. Incipient melting techniques were used to corroborate solidus curves, and X-ray diffraction was employed to study the nitrogen-rich region of the a + ZrN phase field. Materials Westinghouse Grade 1 iodide zirconium crystal bar served as the base material for this investigation. The as-received bars were lightly sandblasted, pickled in a 20 pct HN0-5 pct HF aqueous solution, rinsed in water and acetone, and dried. They were rolled to about 1/32 in. strip and acid-pickled, followed by water and then acetone rinses. The material was sheared to approximately 1/4 in. squares, cleaned with acetone, and stored for use. Pure zirconium nitride is not commercially available, according to correspondence with possible suppliers. A literature survey indicated that nitrogen may be introduced into zirconium metal by passing nitrogen or ammonia gas over zirconium at an elevated temperature. After considerable experimentation, a train was devised whereby high quality nitrogen was passed through a series of bubblers to remove the last traces of oxygen, through an H,SO, bubbler and a cold trap to remove moisture, and finally over zirconium within a resistance furnace. Hand-picked Bureau of Mines magnesium-reduced zirconium sponge proved more amenable to nitriding than the crystal bar. The nitrogen used was extra pure gas purchased from Linde Air Products Co. The first two bubblers through which the gas was passed contained a solution suggested by L. F. Fieser" as being ideal for removing the last traces of oxygen from nitrogen. The solution contains 20 g KOH, 2 g sodium-anthraquinone ,9-sulphonate, and 15 g NaHSO, dissolved in 100 ml water. It is blood-red and turns brown when contaminated with oxygen. A saturated lead acetate solution, next in the series of bubblers, removed any H2S which might have formed in the 0, removal stage. An H,SO, bubbler and cold trap removed moisture before the nitrogen was passed over the zirconium. The zirconium was contained in a stainless steel screen within a Vycor or quartz tube in a resistance furnace. Unreacted nitrogen passed out of the system through a mercury bubbler. A nitriding run was performed in the following manner: Zirconium was placed in the screen within the furnace tube. The unit was assembled, suitably clamped off, and evacuated to remove the air and moisture within the tube and that trapped by the zirconium sponge. The unit was flushed with nitrogen and evacuated twice. After nitrogen was allowed to pass over the zirconium for about 30 min, the furnace was turned on; several hours were required to reach temperature. The reaction was allowed to continue at temperature for about 4 hr and the screen was then withdrawn from the hot zone of the furnace by means of a Nichrome or Kanthal rod. The screen and its contents were allowed to cool at the end of the tube and were then removed. At 800°C no more than about 5 pct N could be introduced into the sponge zirconium. This material was designated M-1 (master alloy No. 1) and reserved for the preparation of alloys in the dilute nitrogen region. A second set of experiments was conducted at 1000°C; no more than 7.2 pct N could be introduced into the zirconium at this temperature. Enough material (M-2) of this nitrogen content was prepared to produce a second set of alloys to complement the first and to provide alloys in the 0 to 6 wt pct N range. It was hoped that a diffusion anneal plus a re-nitriding might yield a material of higher nitrogen content. The material containing about 7 pct N was therefore given a homogenization treatment at 1000°C for 6 hr under an argon atmosphere. It was

Jan 1, 1957