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By L. Adler

Analysis of a mine roof can be based on fixed-end beam behavior. The author here analyzes the effects of zero restraint at deflecting beam supports. Formulae are given for determining permissible support deflections. Analysis of a mine roof in a bedded deposit usually follows the assumption that the behavior of the roof is analogous to a beam with both ends fixed. Yet it is recognized that the end restraint itself varies with the thickness of cover, so that at shallow depth the restraint is minimum, and at great depth, maximum.' Consequently, an analysis on the basis of fixed ends or full restraint represents only one extreme in the possible range of conditions. Since the actual amount of restraint is difficult to ascertain, a more rational approach is to consider the two extreme cases of end restraint. Such situations regarding uncertainty of the restraining effects of supports are not uncommon in structural analysis, and are handled in a like manner.' Since previous studies9)4 have dealt with full restraint, this paper will analyze the effects of zero restraint at deflecting beam supports. In such a case a mine roof with several intermediate supports becomes analogous to a uniformly loaded continuous beam on simple supports, as shown in Fig. 1A. Such a structure is statically indeterminate to a degree equal to the number of intermediate supports, n. Therefore, with the three equilibrium equations, n additional equations are needed. A number of methods exist to analyze this problem, but The Theorem of Three Moments is Simplest. The continuous beam is divided at each support into a sequence of simply supported single span beams, with the unknown end moments, as appears in Fig. 1B. Continuity requires that the common ends of adjacent beams have the same slope (Fig. 1C). eiB= B(i+l)~= ei(i+l)and MiB= Mti+I)A= Mi(i+i). [1I A Glossary of all terms is given in Table I. Furthermore, if we first consider that the supports do not deflect, the slope at a support is the sum of the slopes due to each of the individual loads shown in Fig. 2, yielding:

Jan 1, 1961

By D. Vogt

South African underground hard rock mines are typically managed using measurements made daily, weekly or even monthly of key parameters such as face advance, readiness to blast, blast success, temperature or dust levels. Health, safety and productivity can all be greatly improved if a real-time measurement system can inform decision making. To enable this vision of widespread sensing, communication and decision support, CSIR has developed an open standard architecture for communication of sensor data, and a reference implementation using that standard. The standard, called AziSA, meaning ?to inform? in isiZulu, will allow any device that complies with it to join any network that supports it. The standard is designed for South African gold and platinum mines. To be practical in a typical Bushveld platinum mine, a real-time sensor system needs to be robust, reliable and cheap: ?Sensors are equipped with power sources that will last as long as the sensor is required. In effect, sensors need to be disposable: cheap enough not to warrant repair if they break. ?Communication in the stope is wireless, to avoid there liability problems associated with wire connections. For example, wires to sensors in the face area are unlikely to survive blasting. ?To get information out of the stope, the power line to the scraper winch is used as a communications link. Unlike a separate communications link, the power line will be repaired if damaged, because of its importance to production. ?Outside the stope, communications are transferred onto the mine communication infrastructure. To succeed as a management system, the next key element is the use of software decision support tools that will analyse masses of real-time data to extract key parameters and trends for decision makers. A range of sensors has been developed, including commercial sensors that have had wireless adapters added to make them AziSA compliant, and bespoke sensors that provide low cost sensing of typical mining parameters. An important parameter in any sensor network is the location of the sensor. This has to be easy and cheap to establish, so a tool has been developed to locate sensors in the confines of a stope, using ultrasonic and radio techniques. Once sensors are installed and located, they associate their location to all data that they produce. All the data from the system are captured at surface in a flexible database that can be queried by any application that requires data of any kind. It is simple to program alerts, and for actions to be associated with alerts. The use of wireless in stope also allows for communication from surface to the stope, or from stope to stope. The first implementation of an AziSA compliant system is being installed as part of the Mine Health and Safety Council funded rockfall elimination project. The design is discussed, and a future application for a platinum mine is considered.

Jan 1, 2009

A. G. Cockbain—The paper by Burlingame, Bitsianes and Joseph is of great interest in extending the work done on high grade sinters, particularly that of Hessle, and the development and application to them of the technique of McBriar, Johnson, Andrews and Davies on ironstone sinters. In this laboratory the work of McBriar et al. has also been continued and extended to cover high grade sinters, and a sectioning technique similar to that used by McBriar et al. was attempted on sinter cakes made in the laboratory sinter unit, using magnetite concentrates (containing 64 pct Fe). In these experiments no attempt was made to flush out the products of combustion with nitrogen before cooling. It was considered that free air penetration would occur on only a limited scale with the size of sinter cake made (14 in. sq x 12 in. deep). Chemical analyses of the zones showed generally similar features of those described by Burlingame et al. and especially the FeO rich region at and just behind the flame front, i.e. zones of ignition and combustion. This seems to indicate that the nitrogen atmosphere does not materially affect the states of oxidation in the sinter cake or at least that special care is not required unless the sinter cake is small. Burlingame et al. consider that FeO present has some significant part to play in the mechanism of sintering. This view does not find favor with the author. The reduction of iron oxides in the presence of red hot coke is a well known phenomenon, whether conducted in an atmosphere of nitrogen or in a crucible in air, and in view of the presence of some carbon in the hot sinter, especially just at the ignition zone, cessation of the progress of the flame front will in no way affect the reduction of iron oxides near hot coke. Even in a nitrogen atmosphere some CO and CO, will be generated by reduction and by continual oxidation and reduction of CO as the transfer agent, the oxidation state of the iron oxides can be lowered quite rupidly so long as free carbon exists with enough heat. Knowing the composition of the gas at the flame front (obtainable by probes), it would be possible to calculate the static oxidation state of Fe2O3/Fe3O2 in equilibrium with it and hence obtain a check on the chemical analyses to see if in fact direct reduction of the Fe2O3/Fe2O1 to FeO has taken place. Have the authors attempted this? It appears to this writer that what is required is some means of reducing rapidly the heat content of the ignition zone and immediately behind, and at the same time insuring no oxidation by the air. In ironstone sinters the difficulty of oxidation state of the iron did not arise on account of the very large slag bulk. However, in high grade sinters knowledge of this, and also the mechanism whereby high oxidation states can be obtained in the final sinter, is of great interest. R. D. Burlingame (author's reply)—The question has been raised as to whether the high FeO content found in high-fuel sinters is due to actual reduction at the advancing flame front or due to unavoidable direct reduction during the quenching period. For such direct reduction to have taken place in the freshly-formed sinter, a considerable amount of solid carbon would have had to escape combustion and be available throughout the hot sinter zone during the quenching period. A simple stoichiometric calculation with the limited data available indicates that a zone of freshly-formed sinter over 2 in. wide must have averaged roughly 2 pct C to accomplish the amount of reduction found. Furthermore, the presence of this amount of carbon would indicate that the fuel in the charge had not burned over a narrow front but over a zone at least 2 in. thick. In contrast to such a condition, the data of the present investigation indicate that the combustion of fuel is confined mainly to a narrow front in which high temperatures and reducing conditions favor. the formation of excess ferrous iron.

Jan 1, 1959

By M J. Andrews, P J. White, I Bogie

Active geothermal systems occur widely in Sumatra and in the Taupo Volcanic Zone (TVZ) ofNew Zealand. There are also many occurrences of epithermal gold mineralisation in Sumatra, some of which are economic. Minor sub-economic occurrences of gold have been reported in the TVZ but none have yet been mined. There are both similarities and differences in the nature of the active geothermal systems and mineral deposits in Sumatra and New Zealand, which can be systematically related to the tectonic setting. The thicker crust and consequent predominance of silicic volcanism in New Zealand leads to hydrothermal systems which are little eroded and with limited lateral outflows. The systems in Sumatra are of mixed character, with some of the characteristics of those in andesitic island arcs such as Java or the Philippines, e.g. a higher proportion of systems with identifiable acid magmatic volatiles, a greater potential for deep erosion and lateral outflows than in New Zealand; but modified by the more oblique component of subduction and thickened crust so that some are localised in volcano-tectonic depressions and consequently have high piezometric levels of primary hydrothermal fluid. The unusual occurrence of calcic zeolites in veins associated with gold mineralisation in Sumatra is thought to be indicative of æsatelliteÆ boiling zones along de-gassed hydrothermal outflows.

Jan 1, 1995

By Walter R. Hibbard, George H. Eichelman, William P. Saunders

INTEREST in the various products of the austenite eutectoid transformation in iron-carbon alloys, particularly as produced by the isothermal sub-critical techniques introduced by Davenport and Bain,1 has resulted in the application of similar heat treatments to other eutectoid transformations. Studies2.3.4 of the beta copper-aluminum alloy eutectoid transformation revealed microstructures very similar to those found in iron-carbon alloys. Investigations5,6 of the kappa phase in the copper-silicon alloy system have established the eutectoid type transformation of this phase into terminal alpha and intermediate gamma phases. This transformation has been reported' as forming a eutectoid type structure in a very sluggish manner by a long time isothermal treatment. The purpose of the present investigation is to study the products of the kappa eutectoid transformation as affected by transformation temperatures by means of microscopic and hardness methods. PREPARATION OF ALLOYS The copper-silicon alloy equilibrium diagram is shown in Fig I. The kappa field converges to a minimum silicon solubility at the eutectoid composition and all other completely kappa alloys are hypereutectoid. Thus, kappa decomposition involves first the proeutectoid precipitation of gamma, followed by the eutectoid transformation of a lower silicon content kappa (5.2 pct). Smith5 has indicated that alloys containing between 5.4 and 5.7 pct silicon could be hot rolled, but that the latter composition was desirable in order to permit a wider kappa temperature range. The alloy was prepared by adding Baker's commercially pure silicon to an ingot of copper-silicon alloy containing 2.34 pct silicon remaining from the investigation of Brick, Martin and Angier.7 Following the casting techniques of Krynitsky, Saunders and Stern,8 the ingot was melted under a heavy layer of dried charcoal in a clay-graphite crucible in a gas-fired pot furnace. Small lumps of silicon were held under the melt surface with a graphite rod. After the silicon had dissolved, the alloy was chill-cast and then remelted without a cover, stirred thoroughly, skimmed and poured into a preheated cast iron mold. The cast microstructure is shown in Fig 2. The casting, 10 X 4 X ¾ in., was annealed at 750°C for 21 brand furnace cooled. About 1/16 in. was milled from all surfaces. The plate was cut in half longitudinally and the bottom half analyzed by A.S.T.M. methods.9 The composition by weight was as follows: copper 94.24 pct, silicon 5.56 pct, iron 0.11 pct.

Jan 1, 1948

By Bradley Stoughton

[Secretary's Note.—TO avoid repetition of foot-notes, references to authorities are made in this paper by means of figures, referring to a numbered list in the appendix.—J. S. 1 Introduction. This paper gives the results: first,, of a research into the somewhat extensive literature of titanium; and second, of a series of original tests upon the effect of titanium on iron castr ings. New investigations have also been made upon the effect of titanium on steel, but I hope to treat this subject in an independent paper at some later date, and, therefore, will not discuss the experiments in detail here. The combined results of this study, which extended over a period of more than two years, can be briefly summarized as follows : 1. Steel or cast-iron, in which the titanium is properly proportioned, and which has been properly treated, is improved in strength, toughness, and durability against wear, such 'as, for example, the wear of railroad-rails, of steel and chilled-iron rolls, car-wheels, etc. 2. These improvemerits seem to be caused, not so much by the direct effect of titanium on the metal, as by its cleansing influence in removing harmful impurities, such as oxygen and oxides, nitrogen, occluded slags, and perhaps also sulphur. It appears also to reduce segregation, which would contribute to the same end. 3. In order that its effects may be fully realized, the treatment of steel or iron with titanium must be correctly performed. There are a few simple, but essential, details to be observed; if they are neglected, the best results cannot be expected. To illustrate by an example: A series of tests on cast-iron

Jan 1, 1913

The current understanding of the mechanics of rock fracturing during smooth blasting is largely empirical and qualitative due to an inability to model the dynamic process at the fundamental level. In order to overcome this difficulty, a numerical simulation of smooth blasting is performed using the Dynamic Lattice Network Model. This model is capable of simulating material inhomogeneities and the dynamic evolution of fragmentation at the micromechanical level. In this study, smooth blasting techniques are numerically analyzed to characterize the development of the damage rock zone during explosive fracturing.

Jan 1, 1995

By Lingzhi Yang, Feng Chen, Yufeng Guo, Tao Jiang, Zeqiang Xie

Zinc ferrite (ZnFe2O4), one of the major forms of zinc in Electric Arc Furnace (EAF) dust and zinc leaching residue, is very difficult to chemically dissolve in either acidic or alkaline solutions. In spite of the zinc resources in the Electric Arc Furnace (EAF) dust and zinc leaching residue are abundant, there is no mature utilization technology. This situation not only causes the waste of resources, but also produces the harm to the environment. In the paper, the basic principle of this process has been established by revealing the phase equilibria in the CaO–Fe2O3–ZnO system, where Ca2Fe2O5 can be in equilibrium with ZnO rather than ZnFe2O4 at higher CaO region. It has also been demonstrated that the formed ZnO can be solved by ammonia-leaching and Ca2Fe2O5 keep in solid residue, which can be used as a flux for steel refining.

Mar 1, 2017

By Jeffrey S. Newman

This report establishes a methodology for the evaluation of gas detection devices and is the result of a critical investigation into the basic principles defining gas detector performance. A scheme is presented to evaluate the performance of a test sample of gas detectors based on the detector type and intended application. Test procedures, worksheets, test apparatus, glossary and statistical methods for handling test data arc given. In addition, evaluation standards and test parameters are identified. It is intended that this report serve as a comprehensive framework for the eventual certification of all gas detectors for mine applications. A plan is proposed and recommended for the quantification and implementation of evaluation standards for gas detection device certification.

Jan 1, 1980

By John C. Fowle

Having had for many years the management of magnetite mines, and having noted the various admixtures, such as jasper, " green rock," actinolite, etc., that occur so frequently in the deposits and make the fine ore so difficult as well as expensive to sort by hand, I studied the different ways of mechanical sorting, but found none eatisfactory—the universal objections being the expense, the small quantity that could be handled per hour, and the low grade of ore produced after all. Feeling that some means would be discovered ultimately for concentrating the small ore, I began at the Michigamme mine, at Michigamme, Lake Superior, Michigan, to separate it from the larger and richer lumps, by placing in all the shaft-houses screens of round iron bars set 1/4-inch apart, on which the skips were dumped as they were hoisted; the coarse ore going into one stock-pile or into cars, and the small ore, from 2 or 3 inches diameter down to powder, going into a special stock-pile, to be treated as soon as some proper means was found to separate the ore from the rock and ether impurities. Careful investigations showed that neither dry nor wet concentration was the right process for us; and I finally studied the magnetic process. Not knowing of any magnetic separator in this country that would handle such large pieces of ore as I wished to treat, I looked abroad and found in Sweden the Wenstrom separator. One of these machines was brought to the United States and tested in New Jersey and other places. (See the papers of Messrs. Cook and Birkinbine, Trans., xvii., 599 and 728.) This original separator, having 15-inch face of drum, I brought to the Michigamme mines, and, no one having confidence in it, placed it in operation at my own expense. Finding that in this way large ore could be treated much better by magnetic separators than by hand, and with much greater enrichment of the product, I put up a small, crude mill with several series of screens, so as to size the material and

Jan 1, 1891

By T. Moyo

"The oxidative leaching of chalcopyrite in ammoniacal solutions has been evaluated using electro-analytical techniques and controlled bulk leaching studies. The anodic dissolution process has been established to be a sevenelectron transfer process under nitrogen in ammonia–ammonium sulphate solutions and ammonia–ammonium carbonate solutions, which suggests that the sulphur is oxidized to thiosulphate and the copper and iron released as Cu+ and Fe2+ in these systems. The deportment of Fe2+ and S2O3 2- is affected by choice of the ammonium salt used in the leaching process. In the perchlorate salt, only five electrons are transferred, supporting formation of different sulphur species as proposed for the sulphate and carbonate salts. Scanning electron microscopy and energydispersive spectroscopic analysis of the mineral surface after leaching indicate presence of an iron–sulphur surface layer completely free of copper in the sulphate system, an iron-rich surface layer in the perchlorate solutions, and absence of surface layer build-up in carbonate solutions. XRD analysis of a bulk leach residue from leaching in ammoniaammonium sulphate solutions showed the surface layer to be mostly amorphous (90%). The crystalline content (10%) is composed of 95% polymorphs of anhydrous iron oxide hydroxide FeO(OH). Choice of ammonium salt and the hydrodynamic environment of leaching have been shown to influence the presence or absence of the surface product, as well as its nature. IntroductionHydrometallurgical processes for copper extraction provide a viable route to recover the metal from mixed and low-grade ores, as well as overcome the environmental challenges faced by the traditional pyrometallurgical processes. Chalcopyrite is not only the most abundant of the copper sulphides, but also the most stable, making it recalcitrant to hydrometallurgical processes. Hence, hydrometallurgical processing of chalcopyrite continues to be an attractive area of research due to the vaguely understood surface chemistry of the mineral in different aqueous media. Different routes for hydrometallurgical treatment of chalcopyrite can be followed. These include thermal treatment prior to leaching, direct leaching and direct electrochemical leaching (Venkatachalam, 1991). Direct leaching of chalcopyrite can be carried out in various solution systems, as reviewed by Roman and Benner (1973) and Venkatachalam (1991). Ammoniacal solutions are attractive and effective lixiviants that form stable amine complexes with some base metal cations while rejecting iron. Leaching of chalcopyrite in ammoniacal solutions in the presence of an oxidant is possible due to the stabilization of copper (I) and copper(II) by ammonia at elevated pH levels. In oxygenated ammonia solutions, it has been suggested that chalcopyrite dissolves according to Equation [1] (Beckstead and Miller., 1977a):"

Jan 1, 2016

By Vince Caggiano

The newly developed Hilti OneStep roof bolt combines the standard 5-step anchoring process (i.e. drilling, transfer of drill steel, insertion of resin, insertion of rod and tightening) into one easy operation. It also replaces 4 items (drill head and steel, adhesive cartridge and anchor) for simplified material management. This paper describes its history of development, and bolt, bolt drilling and bolt resin specifications, experimental test results in Germany where it achieved a reduction of 36-42% in roof bolt cycle time and 25-30% improvement in gateroad development rate. In the U.S. coal mines, a time study showed that OneStep can save approximately 50 seconds in installation time per bolt.

Jan 1, 2005

By T. M. Siew

"This paper begins with a brief discussion of the geology of the Newfoundland Zinc Mine area. The results of groundwater studies and pumping tests are then presented. It was decided that groundwater in the mine could best be controlled by conventional sump pumping and grouting. The system is fully described, as well as plans for improvements on the method. LocationThe Newfoundland Zinc Mine is located in the Great Northern Peninsula of Newfoundland, six miles northeast of Daniel's Harbour. The mine is situated at an elevation of approximately 350 feet above sea level, 5 miles farther inland , west of the long range of mountains which rise abruptly with elevations of over 2,000 feet above sea level.GeologyThe ore deposits occur in lower Ordovician carbonate rocks that form part of the stable St. Lawrence platform, which is mainly composed of (1) the lower Cambrian Labrador group, (2) the upper Cambrian to lower Ordovician St. George group, (3) the disconformity overlying Middle Ordovician limestones and shales of the Table Head formation and, (4) two overlying klippen composed of Cambrian and lower Ordovician rocks.The ore deposit is a Mississippi Valley type of the Tennessee variety, having proven are reserves of 5.3 million tons averaging 9.70/0 zinc from eleven ore zones which exist in an area of 5.5 square miles.Among these ore zones, the L zone is the major underground mining area, and is by far the largest of the known deposits. It has proven are reserves of 3.4 million tons with 11.5% zinc; this underground development has been inprogress since 1974.Stratigraphically, the orebody, which occurs about 300 feet below the disconformities between the Table Head and St. George formations, is situated in the upper St. George formationof lower Ordovician age.The upper St. George formation of the L zone is divided into three main units (Fig . I) : the underlying lower limestone the middle grey dolomite pseudo breccia unit and the overlying siliceous dolomite unit. In this area of the mine, the beds of the St. George formation dip gently west to northwest to 10 to 15 degrees. The orebody, which is generally strata controlled with a southwesterly plunge and which has the form of thin, narrow, but elongate, en-echelon ore bed s, is contained in grey dolomite pseudo breccia. Its dimensions vary from 50 to 900 feet in width and 8 to 120 feet in thickness, and the over-alllength exceed s 6,000 feet.The surficial deposit of the area is a till comprising a heterogeneous mixture of sand, gravel, cobbles and boulders, with an average depth of 20 feet."

Jan 1, 1980

By Donald E. Ralston, V. Rajaram, Michael N. Greeley, John W. Peters, Terry J. Laverty, F. S. Kendorski, Peter G. Chamberlain, J. Brent Hiskey, William C. Larson

Preliminary figures from the US Mine Safety and Health Administration show 801 underground metal mines and 490 surface metal mines, at the end of 1981, up from 663 and 383, respectively, in 1980. Preliminary MASH figures also show 207 underground nonmetal mines and 1070 surface nonmetal mines at the end of 1981, compared with 167 and 1092, respectively, last year. In addition, preliminary MASH figures show there were 14,676 mineral extraction operations on the surface and 3845 underground at the end of 1981, versus 14,951 and 3675, respectively, for last year. The number of mines, then, in three of four categories increased significantly, while having a lower net total of just 2% in the other category. And the number of underground extraction operations increased nearly 5% for the year, while surface operations were down less than 2%. But despite operations totals that mostly held their own or showed significant increases, it was a tough year for the nation's mining industry. The nonferrous segment of the industry had a bad year and 1982 may be no better, according to Business Week (Jan. 11, 1982), which adds that a turnaround is impossible unless the general economy recovers. Even then, with diminishing markets and overcapacity plaguing its products, the mining industry's condition "will be far from robust," the magazine says. The basic problem is that nonferrous metals-aluminum, copper, lead, molybdenum, nickel, and zinc-are tied to the most troubled sectors of the US economy. As recently as 1979, for example, more than 40% of all aluminum shipments went to the construction and transportation industries. Last year, these markets were down to 35% of aluminum shipments. And molybdenum sales to steel producers dropped by nearly 15% last year. Moreover, nonferrous products ended 1981 at disastrously low prices. Molybdenum oxide sold at $8.77/kg ($3.98/lb), half its $17.63/kg ($8/lb) price earlier in 1981 and aluminum was listed at $1.08/kg ($0.49/lb), compared with $1.54/kg ($0.70/lb) a year earlier. The Comex price of copper was $1.54/ kg ($0.70/lb), down from $3.11/kg ($1.41/lb) in early 1980, in real terms the worst copper price in 30 years. Lower prices led operators to cut their output and increase layoffs, shut down mines, and defer capital spending projects. In December, Duval Corp., a Pennzoil Co. subsidiary, shut all four of its US copper mines until March. Aluminum Co. of America had laid off 7.3% of its work force, shut nine of its 38 smelting lines, and entered 1982 operating at only 68% of capacity. Inco Ltd., the Free World's largest nickel producer, lost money in 1981, an estimated $11 million, for the first time since 1932. And, like Falconbridge Nickel Mines and other major nickel producers, had cut back on nickel production and was operating at only 63% of capacity (Forbes, Feb. 15, 1982). Amax, the country's largest molybdenum producer, had cut production by 20%. And Hanna Mining Co. sold lead for $5.07/kg ($2.30/lb), about $1.54/kg ($0.70/lb) below what it cost to produce. Another problem is that copper companies in the next few years must come to terms with problems of aging smelters. Some companies are simply closing them. In 1980, Anaconda closed its 75-year-old Butte, MT, smelter and is shipping concentrates to Japan, rather than modifying the smelter to meet environmental requirements. And oth Phelps Dodge Corp. and Asarco Inc. have antiquated smelters on which they will have to make decisions. The most troubled metal in 1982, Business Week said, will be molybdenum, with no real recovery in sight until the mid-1980s. High molybdenum prices in the late 1970s, caused in part by shortages, triggered overexpansion that will take years for the US market to soak up. On a more upbeat note, 1981 mergers might help stabilize the industry by providing funds for mining companies. Standard Oil Co. (Ohio) said it plans to spend $7 billion on Kennecott during the next 10 years for modernization and expansion. And Atlantic Richfield Co. said it will spend $500 million on Anaconda Copper Co. during the next five years. Another positive note was sounded by T. H. Adams of the United Banks of Colorado Inc. He said he expects the mining industry to experience moderate growth in 1982, led by "interest rate sensitive" demand from energy, electronics, and defense markets. And mining will be helped by "weak, but improved" construction and auto markets. The mining industry will be operating in an economy stronger than 1981, but still relatively sluggish by postwar standards, Adams predicted. Iron Ore Estimated US iron ore production in 1981 was 75.2 Mt (74 million long tons), up 6% from 70.7 Mt (69.6 million long tons) in 1980, though down from the 87.1 Mt (85.7 million long tons) produced in 1979. The mine value of usable iron produced from domestic mines was estimated at $3 billion. US iron ore was produced by 28 companies operating 35 mines, 26 concentration plants, and 16 pelletizing plants. The mines included 34 open pits and one underground mine. Byproduct ore recovered from copper- and titanium-mining operations accounted for less than 1% of total iron ore production. Fifteen mines operated by nine companies accounted for 94% of total

Jan 5, 1982

By R. Radhakrishnan, C. Gunasekaran

"Calicut Bypass Phase-II project included construction of 7 m high embankment over soft compressible clays up to 18 m in thickness. Approx 1.5 km length of road out of a total 5 km length consisted of abandoned paddy fields which were low lying, waterlogged and marshy in nature. On the basis of excellent results achieved in earlier phases of the by-pass construction, ground improvement work by installing pre-fabricated vertical drains (PVD) was undertaken to stabilize the embankment foundation along the weak clay areas prior to road pavement construction to avoid large post construction settlement as well as to improve the lateral stability during and after embankment construction. This project perhaps stands out as one of the highway projects completed within a very short time period of total 18 months including four month long heavy monsoon period, when embankment construction was not possible and construction of two bridges across rivers. All activities including the ground improvement had to be programmed and completed within the shortest time period possible. The paper discusses details of the extensive ground improvement and soil instrumentation carried out for the project and will illustrate how careful planning and execution can reduce the period required for highway projects involving ground improvement.INTRODUCTIONPhase-II construction of Calicut by-pass was undertaken on a priority basis by the State of Kerala in 2014. The main contract was awarded to M/S ULCCS Ltd, Kozhikode, Kerala and the ground improvement work was awarded to M/S Geo-Enviro Engineers P Ltd, as specialist agency to design and execute the work.Soil investigation carried out confirmed thick soft clay deposits below GL in three areas along the road alignment totaling approx 1.5 km in length. These areas were low lying, waterlogged and marshy in nature. Highway embankment had to be constructed up to 7 m in height above existing GL to form the proposed road alignment along these areas. On the basis of excellent results achieved in earlier phases of the by-pass construction using pre-fabricated vertical drains or PVD (R. Radhakrishnan, 2006), ground improvement using PVD had been included in the contract to stabilize the embankment foundation along the weak clay areas prior to highway embankment construction to avoid large post construction settlement as well as to improve lateral stability during its construction."

Jan 1, 2015

By R. C. Earlougher, F. G. Miller, T. D. Mueller, H. J. Ramey

There are many studies of flow in radial systems that can be used to interpret unsteady rerervoir flow problems. Although solutions for systems of infinite extent can be used to generate solutions fur finite ow systems by superposition, application is tedious. In this paper a step is made toward simplifying calcu1ations of such solutions for finite flow systems. Superposition is used to produce a tabulu-lion of the dimension1e.s.s pressure drop function at several locations within a bounded square that has a well at its center. The square system provides a useful building block that may be used to generate flow behavior for any rectangular shape whose sides are in integral ratios. Values of the tabulated dimensionless pressure drop function are simply added to obtain the dimensionless pressure drop function for the desired rectangulm system. The rectangular system may contain any number of wells producing at any rates. Furthermore, the outer boundaries of the rectangular system may be closed (no-flow) or they may he at constant pressure. Mixed conditions also may be conyidered. Tables of the dimensionless pressure drop function for the square system are presented and various applications of the technique are illustrated. Introduction In 1945, van Everdingen and Hurst' published solutions for the problem of water influx into a cylindrical reservoir. Since this problem is mathematically identical with the depletion of a cylindrical reservoir with a well at the origin, the van Everdingen-Hurst solution may be used to study the depletion problem. In their analysis, they assumed that the fluid had a small, constant compressibility such that flow was governed by the diffusivity equation For a constant production rate q starting at time zero, van Everdingen and Hurst showed that the unsteady pressurc distribution for both finite and infinite systems could be expressed in terms of a dimensionless pressure 'Sabulations of the dimensionless pressure drop for a unit value of r, were provided by van Everdingen and Hurst,' and later by Chatas.' Others also presented values in graphical or tabular form." If the radius of the well becomes vanishingly small, r,+ O, the line source solution may be used for Eq. 2 when infinite systems are considered. Eqs. 5 and 6 are excellent approximations for Eq. 2 under certain conditions: In 1954, Matthews, Brons and HazebroekV emonstrated that solutions such as Eq. 5 can be superposed to generate the behavior of bounded geometric shapes; i.e., the behavior of a bounded single-well system can be calculated by adding together the pressure disturbances caused by the appropriate array of an infinite number of wells producing from an infinite system. These wells are referred to as image wells. Matthews, Brons and Hazebroek considered systems containing a single well producing at a constant rate. This superposition can be represented analytically as

Jan 1, 1969

By David Ballón Vera

El presente estudio describe la importancia del financiamiento en la industria minera como herramienta de difusión a través de dos condiciones: la primera, es que esta se adecúe a las características de la minería y la segunda, refiere que los mineros replanteen su estructura organizativa de estudios (pre-factibilidad,factibilidad, ingeniería y diseño) para el inicio de una posible inversión, así como preveer que de la experiencia obtenida, se contó con poco apoyo financiero.

Nov 20, 1978

By Arnulf Muan, Klaus Schwerdtfeger, L. S. Darken

The equilibrium ratios CO/CO, of. a gas pharse coexislitrg with selected oxide phase assemblages of the system "FeO"-MnO-SiO2 and metallic iron have been determined at 1150°C. The data obtained are combined wilh the thermodynamic data on fayalite (Fe2SiOi) and "FeO"-MnO solid solutions to derive the actiuity-coiHpQsiUon relations for the various oxide components of the olivine (Fe2SiO4-,lIn2Si04) and pyroxenoid (standard free energies of formation of ferrosilile (FeSiO,), tephroite , and rhodonite from their oxides have been evaluated. In the Appendicc~s by Darken and Schivcerdtjeger the ideal solution law for solid deri oed. THE quaternary system Fe-Mn-Si-0 is of interest in many important phases of iron- and steelmaking. For instance, reduction in the blast furnace of ores containing Fe-Mn silicates, deoxidation of liquid steel with Mn-Si alloys, formation and subsequent alterations of oxide inclusions in steels, and the oxidation of manganese- and silicon-containing steels during processing are all governed largely by the thermodynamic relations existing in the system Fe-Mn-Si-O. Although the equilibria between liquid iron and Fe-Mn silicates have been studied in detail previously,' experimental data are lacking for equilibria among the solid phases. The present paper provides such data for a selected temperature, 1150°C. I) PREVIOUS WORK The general feature of subsolidus phase relations in the system iron oxide-manganese oxide-silica under strongly reducing conditions is known.' Fig. 1 represents an isothermal section at 1150°C. The phases present, in addition to silica, are three solid-solution series, viz.. the oxide solid-solution series "Fe0"-MnO (manganowiistite), the orthosilicate solid-solution series Fe2SiO4-Mn2SiO4 (olivine), and the metasilicate solid-solution series FeSi03-MnSi03 (pyroxenoid). The "FeO"-MnO solid solution is continuous between the two end members wustite and manganositeV3 as is the orthosilicate series between the two end members Fe2SiO4 and Mn2 SiO4. 4 The metasilicate solid solution, on the other hand, extends only part way from the MnSi0.l end member toward FeSiO because the latter end member is unstable relative to the phase assemblage fayalite + silica. The letters a through m in Fig. 1 identify end members of the individual solid phases dealt with [for instance, c designates fayalite (FeZSiO4) in equilibrium with wustite, e designates fayalite in equilibrium with silica, and so forth]. In addition to the phase relations there are some thermodynamic data available for the end members of the silicate solid solutions and for the "FeO" -MnO solid solution. The free energy of formation of fayalite (Fe2Si04) from its oxide components has been studied by several investigators. Unfortunately, the data obtained show considerable discrepancies. According to recent tabulations by Kelley, the change in standard free energy Affn for the reaction

Jan 1, 1967

By George A. Koenig

This new departure in assaying is the outcome of a long-felt desire to shorten the time required in muffle-assaying, as well as to do both crucible- and scarifi cation-work in one furnace. The first object in view was the gold-assay, which must always be a combination of the crucible- and muffle-work. The large lead buttons consume much time in their reduction by scorifying and cupelling. The Hoskins gasoline furnaces have been a great convenience, for years past, in that class of practice, where only one assay was required at a time. But even with them I found myself hampered, since the crucible-furnace only takes one crucible at a time, and my practice implies always duplicate assays. Another difficulty was to get the muffle up to the necessary heat and to maintain it at that heat. Whecever I had pumped myself into perspiration and out of breath, the question presented itself, " Is the muffle necessary to do this work properly ?" Hartmann and Plattner and all their blow-piping disciples have shown that very good results are obtainable by playing a direct flame upon the work-lead. Could this be done on the regular muffle-assay scale ? I placed a cupel upon a fire-brick, surrounded it with charcoal, and then played upon it with a Bunsen gas blow-pipe. The result was encouraging. With some care a 2.75-inch scorifier charged with 0.1 A. T. of a 100-ounce ore could be melted down and scorified without any apparent loss. I thought then of constructing a furnace with four blow-pipes in a row; but the combination of gas and charcoal seemed unpractical. Instead, I placed an F crucible, bottom up, in a Hoskins crucible-furnace. Upon this support I placed a charged scorifier. The rapidity with which a very refractory charge melted and scorified was simply amazing to me. Then a cupel was put in the place of the scorifier, the lid was moved sideways, and a blow-pipe was

Jan 1, 1899

There is an increasing demand for refrigeration as mines progress deeper. Mine refrigeration plants have special requirements in terms of their duty and application. There are no national or international standards applicable for the testing of these plants. There have been numerous examples of plants failing to meet their required nominal capacity, but with the plant owner being unable to enforce the design duty on the manufacturer, due to a failure to unambiguously specify testing or accepting criteria. This paper discusses some of the reasons for performance shortcomings and describes a methodology for developing and specifying performance criteria, when inviting design and/or construction tenders, and a method of performance and acceptance testing once the plant has been built and commissioned. These guidelines encompass the full range of performance duties expected from a mine plant, including stable operation at maximum and minimum plant capacity, and under the various combinations of extremes of high and low condensing and evaporating pressures. A description of a methodology for developing a computer-based mine refrigeration plant model is also presented as part of the performance testing criteria. The methodology is illustrated with reference to the actual performance and acceptance testing process used on the worldÆs largest bulk air cooler (R67) at the Enterprise mine at Mount Isa, Australia.

Jan 1, 2005