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By William L. Meyer

With high expectations and promises, computers were introduced to the mining industry in the late 1950s and 1960s. Applications spread in the 1970s when capabilities began to catch up to early promises, and people within the industry began to see the benefits of using computers in business and engineering. The 1980s have seen a wide acceptance of computers, both large and small, and the use of computers is accelerating throughout the industry. Computers have changed dramatically over the last 25 years. The power of some microcomputers matches the capabilities of main frame computers from less than 10 years ago. The variety of software on these micros exceeds what is available on many larger computers today. Changes over the next 10 years are likely to be greater than those of the last 10 years. Making use of available computer technology while trying to cope with the rapid changes in this industry will take much effort within mining. To take advantage of existing computers, consider and evaluate: project needs and applications; professional, administrative, and support personnel; computer software and technology; and computer and related hardware. Note that this list is in the order of decreasing importance. These categories are also presented in the order for which organization and solutions should be provided to get the most out of computers. To start with a computer and go up the list has cost many mining companies needless expenditures, wear and tear on personnel, and frustration. It is assumed for this article that the need for computer tools is self evident. You can do without a computer by using the appropriate techniques (manual methods) with any required shortcuts plus a lot of personnel time. Indeed, several companies are doing that as a "solution" to what to do with their excess engineering personnel in these lean times. However, the future may be better served if some of these people could be prepared to handle the applications and technology presently available. People and Problems To present the factors related to turning the computer into a productivity aide, it may be enlightening to examine some common people problems that reduce productivity. The most valuable resource any mining company has is its mining professionals. And, particularly with computers, people either make the successes or create the barriers, not the machines. The following list of people problems (syndromes and complexes) has been amply observed during some 23 years of the author's experience with computers and their applications in mining. Home Grown Syndrome - "Nobody else can really understand my problem or provide `exactly' what I need to solve it." While there may be some truth to this syndrome, it would be an unusual project that cannot be handled with existing solutions for at least 75-80% of the needs. Having a base to build on will also speed the development of any remaining needs. It does not seem justified to start all over for a new application, even if "our problem is unique." Not Invented Here Syndrome - This "justification" for the continued existence and budgets of some groups has various forms: "It wasn't done here so it obviously can't do the project the way I think it should be done"; and, "If it wasn't invented here, it obviously

Jan 11, 1983

By D. Pearson

About 15,000 tonnes of zinc per year is present in the dusts and fumes produced in electric arc and BOS steelmaking, secondary copper refining and brass ingot making. Apart from the use of some copper-smelter fumes for zinc chemicals manufacture, zinc is not recovered and their disposal by landfill entails some risk of pollution by leach-out. The paper describes processes for recovering the zinc based on leaching and electrowinning. It was shown that the procedures normally employed for purifying zinc-electrolytes derived from primary sources are sometimes ineffective for secondary materials. Alternative process routes are proposed. Source materials were leached with acidic or alkaline reagents and solvent extraction was employed for electrolyte purification in addition to the usual techniques involving precipitation. Examples of integrated flow sheets for specific source materials are presented.

Jan 1, 1980

By Ronald R. Bauer

Wollastonite is a little known industrial mineral of the silicate group. The largest and purest known deposit in the free world is being mined and beneficiated in Millsboro, New York by NYCO, A Division of Processed Minerals Incorporated. The associated garnet and diopside are separated from the wollastonite using dry high intensity magnetic separators. Beneficiated wollastonite is used extensively as an additive and filler in the ceramic, paint, refractory, and wallboard industries. Wollastonite acts as a functional filler due to its unique chemistry and physical acicularity.

Jan 1, 1982

By Robert E. Lovett

The truly educated man is better prepared to meet the challenge of the times. This challenge can be either beneficial or harmful depending upon one s orientation, Certainly man can no longer continue in the way of his father or even in the manner of his own previous experience as he seeks self-fulfillment, unless he is willing to sacrifice inner integrity--unless he is willing to compromise self-determination and self-expression. He must make a choice as society moves rapidly into high gear in the areas of automation and cybernetics. Industry, a unit of society, has a high stake in meeting this challenge, The better prepared and more mature its personnel, the better the end result as it attempts to meet the challenge. It therefore has a vital concern in proper training of such personnel, that their contribution will be of maximum quality. The specialized training required demands a broad liberal arts background which gives the individual a wider perspective upon which to base judgment. This greatly aids his development as he enters graduate school or goes directly into industry. In either case, his value to industry is enhanced. It would seem that the small liberal arts college provides the best environment for such a program, Consideration of several factors will, I hope, throw some light on the severity of the problem.

Jan 1, 1966

By G. M. Wilemon

The U.S. Bureau of Mines has conducted preliminary investigations on the leaching of phosphate ores obtained from deposits in the western United States using sulfuric acid (H2SO4) in the presence of methanol. Phosphate extractions in excess of 80 pct were obtained on an ore sample from Pocatello, ID. Initial leaching studies using the ore as received yielded extraction levels of approximately 35 pct. However, when the sample was ground for as little as 5 min in a rod mill, in methanol, then treated with HzS04, 82 pct extraction was achieved. The R2O3-P2O5 (R2O3 = Al2O3 + Fe2O3) ratio of the starting material was 0.11. Solubilized products after leaching exhibited an R2O3-P2O5 ratio of 0.02, which is well within desired limits. These results illustrate the potential of acid-alcohol leaching for recovering phosphate values from marginal-grade ores and as an alternative to methods involving current beneficiation techniques that yield lower phosphate recoveries.

Jan 1, 1992

By W. C. Brice

A cooperative study was undertaken by representatives of the environmental community, the mining industry, and state government to identify and resolve environmental issues associated with base and precious metal mining in a neutral atmosphere before a commercial mining development is announced. The study included testing Minnesota's as-yet-untried non-ferrous mining regulatory program using three hypothetical mining developments sited in environ-mentally sensitive areas where future mining could occur. Results include identification of critical paths for regulatory decision making, identification of particularly sensitive environ-mental issues that will need continued deliberation before resolution, and characterization of data necessary for environmental review and permitting decisions.

Jan 1, 1990

By Doug Williams, Kent J. McGrew, Dr. Jack Murphy

For some time, selenate has eluded water treatment by conven­tional means such as precipitation and clarification. This has been due to the nearly universal solubility of all oxidized sele­nium species, the low levels required for surface discharge, and almost identical chemical behavior to all forms of oxidized sul­fur. The last characteristic makes selenate particularly difficult to deal with in acid mine waters where sulfate can easily be on the order of one gram per liter. Approaches such as ion exchange or reverse osmosis often tum into total demineraliza­tion schemes for the sake of removing a "trace" selenium species that may be present on the order of 108 sulfate/selenate. Reduction of selenate to selenite or elemental selenium by chemical means has been relatively unresearched. Some limit­ed success has been reported (Murphy, 1988) using ferrous hydroxide as a reducing agent; however, the reaction time as reported is too slow to be useful in a practical plant design. One significant difference between selenate and sulfate is that selenate has been discovered to be reducible in acid solution by elemental iron when catalyzed by soluble copper. Since the +4 or zero valence states of selenium are rather easy to deal with by coprecipitation with iron salts or alum, the elemental iron dis­solved in the process contributes to the overall process by pro­viding sludge upon raising the pH that can be separated by con­ventional clarification. Commercial production of selenium using copper catalyzed reduction of selenite to selenium metal via sulfur dioxide is well established, however, at the < 5 ppb requirements for surface discharge this reduction process equili­brates at levels far too high to be useful to low level selenate removal criteria. The reaction reported herein uses elemental iron for reduction of selenate with copper acting either as a cata­lyst or as a coprecipitation agent. This method, employed with conventional sludging, clarification and filtration can provide an economic method to achieve non detectable selenium (< 0.005 mg/I). Iron consumptions for the conditions tested have been on the order of 3 lbs/1,000 gallons treated. This represents as little as $0.06 or $2.00 per 1,000 gallons based on using a scrap source of iron or the purchase of electrolytic iron for the process.

Jan 1, 1996

By Roy A. Grancher

Construction materials indeed face a dilemma...as does the building industry as a whole. A stop & go economy; a severe drop; in housing starts; lingering indecision over just what corrective steps are to be taken or to their degree of commitment; no less basic disagreement in regard to which of the, Nation&apos;s problems is really the most pressing -- inflation or recession? Optimism has become a rare quality these days. At moat the cautious view is that this new year will see a long, hard struggle back, with hopes that 1976 at least can truly mark a dual anniversary celebration for the United States and its economic health. Yet the paradox is that once recovery does return, those very construction minerals and materials --which in recent years have experienced the ups and downs of overcapacity, shortage, and now declining demand again-- will once more face the spectre of mass inability to service revived and growing needs for its product. Mounting costs, present marginal expansion investment, and substantial lead times required for new facilities to come into production, could well form a bottleneck to economic upswing. The result again can be another cycle of shortage-induced inflation, and with it continued gyrations within the demand supply relationships of the country&apos;s building industries.

Jan 1, 1975

By Rees D. Groves

Cement copper is ordinarily converted to marketable metal by processing in a conventional smelter and refinery. Because of rapidly increasing cement copper production and limited excess smelting capacity, other techniques were investigated for preparing marketable copper products. One route was to make copper powder for industrial use. The process scheme employed flotation for rejecting the bulk of the impurities, acid leaching to remove residual iron and aluminum, reduction sintering to eliminate oxygen, and grinding to obtain powder. Another route was to sinter the floated and leached copper into the form of anodes and electrolyze these to yield refined copper cathodes. At a current density of 30 amperes per square foot, power consumption for electrorefining was 185 kwhr per ton of copper. The feasibility of preparing commercial-size anodes by sintering is still to be assessed.

Jan 1, 1970

By Gerald W. Elger

Research at the Albany Metallurgy Research Center has shown that high-purity rutile can be produced from relatively low-grade ilmenite obtained from a massive rock deposit. Ilmenite is first smelted in an electric-arc furnace to separate the iron and produce a slag enriched in titanium dioxide. Titania slag mixed with a phosphate flux is then oxidized to synthesize rutile and form a phosphate-glass matrix which contains the impurities. Attrition scrubbing the treated slag in dilute solutions of phosphoric or sulfuric acid liberates the rutile, which is concentrated by physical methods. Tests show that synthesized rutile containing 95 to 97 pct titania responds to chlorination at the same rate as natural rutile.

Jan 1, 1973

By F. W. Petersen

Slurry reactors are being used extensively throughout the chemical, biochemical and mineral processing industries. This paper describes the interaction between fine ore particles and granular porous adsorbents such as activated carbon and ion-exchange resins in a stirred tank reactor. Gold and silver cyanides were used as adsorbates to detect changes in the mass transfer characteristics of the adsorbents, which were ashed to determine the loading of ore particles. A simple kinetic model incorporating film diffusion, intrapanicle diffusion ~d the adsorption equilibrium, was used to interpret kinetic data and to estimate the effect of slurry conditions on the diffusion parameters. It was observed that the equilibrium loading of fines in the adsorbents was independent of their concentration in the slurry. Furthermore, the presence of these panicles inside an adsorbent decreases the effective diffusion of gold or silver cyanides into the adsorbent matrix. When fine suspended solids are present in the reactor, a decrease in the effective availability of the superficial surface of the pre-saturated adsorbent panicles for mass transfer from the liquid phase was observed. This newly defined surface availability factor was dependent on the concentration, panicle size, and the zeta potential of the fine particles in suspension.

Jan 1, 1993

By Alphonse C. Van Besien

It is doubtful if anyone knows precisely how many U.S. coal miners have been killed by roof falls. In 1931, 51 miners were killed by falls or rock in every 10-day period (Paul, 1). Roley (2) indicated that from 1882 to 1945, over 3,700 men were killed by roof falls in Illinois alone. During this 63-year period, roof falls accounted for 47.6 percent of the coal mine accident fatalities. Similarly, Price and Nolting (3) reported that during the 10-year period beginning in 1937, roof falls claimed over 50 percent of the lives lost in West Virginia coal mines. Over 150 miners died under fallen rock in West Virginia alone, in each of those years. Certainly, the statistics have improved over the years. The Theodore Barry Report of 1971 (4) indicated that during, the previous 4 years, fewer than 100 miners per year lost their lives to roof falls--in the entire United States. The improvement is less remarkable when it is realized that the accident rate has declined only as the number of miners employed underground has declined. The TRW Report (4) pointed out that although the number of fatalities in underground coal mining (from all causes) had been cut in half from 1956 to 1965, the fatality rate per thousand miners had remained virtually constant. The decreased accident rate was due not to a reduction in hazards but to a reduction in work force. Roof falls still claim approximately 50 percent of the lives lost in underground coal mines. Considering the increased underground production of coal which must result from our energy crisis, the environmentalists&apos; pressure against strip mining, and our inevitably diminishing supply of strippable coal, it is apparent that some means must be found to reduce the number of lives claimed by coal mine roof falls.

Jan 1, 1973

By Heinz Petry

In deep strip mines the job of moving overburden may be so much bigger, and more costly than the removal of coal or ore that the operation is often called dirt mining. Constant improvement in stripping machines and techniques of using them is needed to offset rising costs and to increase production in mines that are now operating, and to make it economically possible to work deeper or thinner deposits.

Jan 1, 1962

By C. E. Davis

The Jamaica Mining Legislation requires the bauxite companies operating in the island to reclaim all mined-out lands (except those legitimately excluded) to the level of agricultural or pastoral productivity, which existed prior to the commencement of mining. If it is not possible to reclaim a mined-out area the companies are required to reclaim an equivalent area of some other land which had not previously been used for agricultural purposes, or alternatively, pay to the government the sum of $120 (US) for every acre of land mined in the particular sector. In the seventeen years since the programme commenced the companies have in general chosen to reclaim most of the mined-out lands, rather than the cheaper alternative. However, the ultimate success of the programme will depend not only on actual reclamation per se but moreso the productivity of the rehabilitated lands. There are a number of reasons why the accent has to be placed on productivity of the lands. The first is that Jamaica is a small country (4,450 square miles) with a population of nearly 2 million people. Because of its mountainous terrain it is estimated that only 35% of the land area is cultivable in crops and pastures; hence this limited area, as well as that which can be reclaimed, must be made highly productive in order to support local and export demands. The Mining companies own over 200,000 acres of land, and of this acreage, it is conservatively estimated that about 75,000 will be affected by mining and will be reclaimed. As such, the use of the land after mining can have an important bearing on Jamaica&apos;s future agricultural productivity. The second reason is that a significant amount of the land now owned by the companies was purchased from thousands of small

Jan 1, 1972

By Lance Rasband, Richard Hanke, John Starcevich

"Mining on Seattle’s SR99 tunnel began in July of 2013. Initial mining progressed slowly through a 4.5 m (14.9-ft) thick fiber-reinforced concrete headwall before immediately entering the first of three planned safe havens. Once through the headwall and safe haven the machine would move through roughly 122 lineal m (400 ft) of jet grout improved soil at a rate of slightly more than 1.6 m/d (5.3 ftpd). It was anticipated that once mining has passed through the improved area, “Bertha” – now the world’s second largest tunnel boring machine (TBM), would be moving at a rate of almost 11 m/d (36 ftpd). By the end of October 2013, Bertha had mined roughly 131 m (430 ft), averaging 1.4 m/d (4.7 ftpd), and had begun mining within native soils. By the start of December 2013, tunneling had reached the 244 m (800 ft) mark and Bertha was averaging 3.6 m/d (11.9 ftpd), achieving as much as 12.8 m (42 ft) of progress on given days. Unfortunately, on Dec. 6th, 2013, Bertha began to stumble and all progress quickly came to a halt.Initial reports stated an obstruction had stopped Bertha in her tracks. The preliminary investigation involved installation of 10 deep dewatering wells to reduce the hydrostatic pressure within the ground around the TBM, to allow the contractor’s personnel to inspect the cutter head from within and also to investigate for the presence of obstructions at the face of the cutter head. During inspection, the contractor’s crews found fragments of steel pipe entangled in the cutter head. The steel was apparently from an abandoned 254 mm (0.9 in.) diameter monitoring well casing, previously installed by the Washington State Department of Transportation (WSDOT) to monitor groundwater conditions for the design and planning of this tunnel project. After the steel was removed from the cutter head and cleared from Bertha’s path, she attempted to move forward but was only able to generate a slow crawl."

Jun 1, 2016

By J. C. Irvine

Meramec Mining Company, a joint venture by Bethlehem Steel Corporation and St. Joe Minerals Corporation, is engaged in the mining, concentrating, and pelletizing of iron ore from the Pea Ridge Mine located near Sullivan, Missouri, about 70 miles southwest of St. Louis. The orebody was delineated in the mid-1950&apos;s by St. Joe during an exploration program to discover new lead orebodies. An airborne magnetometer survey which was run by the U. S. Geological Survey, the Missouri Geological Survey, and several min¬ing companies was published in 1950 and showed an anomaly in the Pea Ridge area. Test holes drilled in 1953 and 1954 revealed the presence of a large orebody of probable commercial significance; additional drilling in 1956 and 1957 provided confirmation of this. Meramec Mining Company was incorporated in 1957 and shaft sinking began in the latter part of that year. The Pea Ridge orebody lies under about 1300 feet of sediments. It is tabular, about 2600 feet long and up to 600 feet thick, and lies in an approximate east-west delineation. The orebody dips about 800S and is of unknown depth. The ore is primarily high grade magnetite with smaller zones of specular hematite, The Pre-cambrian wall rock is rhyolite porphyry. The mine is serviced through two concrete-lined shafts, 19 feet 1 inch in diameter and 2500 feet deep. No. 1 shaft, equipped with an "A" type head-

Jan 1, 1972

By J. W. Francis

The paper about to be presented deals with the relatively straightforward problem of mining two seams of coal separated by an interval. This may seem to be a simple problem but within a twenty-mile radius of Consolidation Coal Company&apos;s Burning Star # 4 Mine at least seven (7) different techniques of two-seam mining have either been tried or are now in use. This great variety of techniques should point out that there is no single simple solution; a conclusion often reached when dealing with mining problems. As a background for the techniques presently in use at the Burning Star # 4 Mine, we will briefly outline the other five (5) stripping techniques used in the area: 1. Captain Hine 6360 Range Diagram (Figure 1) 2. Leahy Mine Range Diagram (Figure 2) 3. Burning Star # 3 Mine 5860 Range Diagram (Figure 3) 4. Burning Star # 3 and # 4 Mines Dragline in Pit Range Diagram (Figure 4) 5. Captain Hine 2570/5561 Range Diagram (Figure 5) Each of these techniques was chosen because it provided the best and least expensive alternative for mining the coal under the particular restraints encountered. The Mine Engineers at our Burning Star # 4 Mine, Mr. Dennis Mahan and Mr. William Weaver, will now detail two additional mining techniques, out-lining our experience to date.

Jan 1, 1977

By W. Jacobi

INTRODUCTION The radon problem concerns the radiation exposure from inhaled short-lived radon daughters and the associated risk for the induction or promotion of lung cancer. In several groups of miners which were exposed in the past to relative high radon levels during their underground work, an excess lung cancer frequency has been observed. On the basis of these findings it has to be stated that the inhalation of radon daughters is the most important radiation hazard in mining. Involved in this radon problem in mines are the difficulties to measure the individual radon daughter exposure of miners and to estimate the dose to critical cells and target tissues in the lung from inhaled radon daughters. And finally the reduction of radon levels in mines confronts us with special technical problems which are difficult to solve. Thus the setting of annual limits for the radon daughter exposure of miners alone is not sufficient to guarantee an appropriate standard of radiation safety. Of same importance are the improvement of air monitoring and the optimization of technical protection measures in mines. This has to be kept in mind for the following considerations on exposure limits. The ICRP has recommended in 1959 for the first time a maximum permissible concentration (MPC) for the occupational exposure to radon and its short-lived daughters (ICRP, 1959). This MPC in air was derived from a maximum permissible dose equivalent of 0.15 Sv (15 rem) per year to the epithelium of the large bronchi which was regarded as the "critical tissue". This old MPC-value was derived using a simplified lung model which paid main attention to the dose contribution from the inhalation of unattached 218Po(RaA) atoms. Since this time more realistic dosimetric models for radon daughters have been developed; some of these models will be presented in special papers on this conference. Furthermore in the last 10 - 20 years substantial epidemiological information on the lung cancer risk among several groups of Rn-exposed miners has become available. Taking into account these epidemiological and new dosimetric findings the ICRP has adopted in March this year new recommendations on occupational limits for inhalation of 222Rn, 22ORn and theirshort-lived daughters. This report will be published this year as ICRP Publication 32 (ICRP 1981). The recommended limits for radon daughters were derived from the basic dose and risk limits as they were proposed by ICRP in its new basic recommendations (ICRP Publ. 26, 1977). In the following this basic system of dose limitation is shortly outlined before the special recommendations for radon daughters are described. [BASIC CONCEPTS AND LIMITS] The new basic recommendation of ICRP (Publ. 26, 1977) proceed from the assumption that in the dose range which is relevant for radiation protection, the risk RT of so-called stochastic radiation effects (carcinogenic and genetic effects) in a target tissue T increases proportionally with the mean dose equivalent HT in this tissue: RT = aT HT. If several target tissues are irradiated the total individual (stochastic) risk is given by the sum of risks in these tissues: [R = ?TRT = ? T aT HT]. Thus this concept of individual health detriment takes into account the risk contribution from each relevant tissue. It replaces the old concept of the critical tissue, which neglected the exposure of other tissues. Applying this new concept a dose quantity can be defined which is proportional to the total individual radiation risk from stochastic effects. This so-called "effective dose equivalent (HE)" is given by the sum of the weighted dose equivalents in the relevant tissues: [HE = E wT HT with E wT = 1 T &apos;1] The weighting factor wT of each tissues follows to wT = aT/awB, where awB is the total stochastic risk coefficient for a uniform irradiation of the whole body. For the lung as a single organ comprising several target tissues, the ICRP recommends a total weighting factor w = 0.12 (ICRP 1977/79). This concept of risk or detriment, respectively, has important consequences. To meet the objectives of radiation protection the new ICRP-system of dose limitation is based on the following three principles: 1) The justification of practices involving radiation exposures; no pratice shall be adopted unless it produces a net benefit. 2) The ALARA-principle that all exposures shall be kept as low as reasonably achievable; this requires the optimization of the level of protection. 3) The limitation of individual exposure (annual dose limits). It is important to note that the dose limits for individuals are only one part of this system of dose limitation. Furthermore, these limits cannot be regarded as a boundary between safe and unsafe. They have to be interpreted as the lower boundary of an unacceptable exposure region. This means that in addition to the observance of these basic dose limits it is demanded to optimize the level of protection. Consequently the optimized exposure levels or the corresponding operational limits will be normally lower than

Jan 1, 1981

By Augusto Andres Torres Toledo, João Felipe Coimbra Leite Costa, Luciano Nunes Capponi, Diego Machado Marques

Short-term mine planning to open pit requires defining diglines to orientate the excavation geometry. In deposits with multiple chemical variables, it is necessary to have more rigorous control over the feeding of the processing plant. This study analyzes the impact of incorporating multiple chemical variables, the risks generated by geological uncertainty, and the number and starting point locations of shovels in the sequential design of diglines. The objective is to maintain and preserve the low-grade variability, required tonnage, and continuity in sequentially selecting blocks. To achieve this, the probability distribution function of each chemical variable is used while attempting to maintain the ore grades as constant as possible. As it is difficult to maintain multiple elements with low variability simultaneously, the relevance of each chemical variable in the economic benefit is identified. Weights are assigned to signal the importance of each element based on its significance in the recovery process. Several diglines, consisting of sequential blocks, were generated on the same seed point to meet the weekly ore demand. Optimizing the stationarity of multiple chemical variables, accessing and reducing the geological uncertainty performed through geostatistical simulation of grades and transfers functions, considering the number and location of shovels, while meeting the operational restrictions of the selected blocks. The optimized digline with the multiple grades as stationary as possible is retained. The process is illustrated in a phosphate mine. The results show that the proposed method is viable and allows for a quick solution.

Jun 25, 2023

By R. A. Pyper

Laboratory leaching experiments were conducted to ascertain the effectiveness of acidic solutions of thiourea, CS(NH2)2, for dissolving gold from finely disseminated gold ores. The ore tested was from the Chipmunk orebody of the Northumberland mining district in Nye County, Nevada, USA. Extractions of up to 90 percent were attained from the ore. Extraction was shown to be dependent on thiourea concentration, ferric sulfate oxidant concentration, acid concentration, and leaching time. Optimum conditions were an ambient temperature leach near pH 1.0 utilizing 100-200 kg/t thiourea and 50 kg/t ferric sulfate for 8 to 24 hours.

Jan 1, 1981