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By Gary W. LaChapelle

Cyprus Miami Mining Corp. has a history of overcoming adversity through innovation and adaptation to changing conditions. Requirements to remain profitable in the 1990s include a combination of advanced technology, employee involvement and training, attention to customer satisfaction, environmental awareness and quality management. The Miami property is located in the Globe-Miami area of Arizona, 130 km (80 miles) east of Phoenix and 160 km (100 miles) north of Tucson. The complex is located in a high desert environ¬ment at an elevation of 1065 m (3500 ft). The region receives about 480 mm (19 in.) of precipitation annually. History Mineralization at the Miami property was first discovered in the 1880s, which led to underground mining of copper by the 1890s. In 1911, the Inspiration Consolidated Copper Co. was formed by combining the Inspiration and Live Oak Mines. Full-scale production began in 1915. Block caving methods were used at both mines. Open-pit mining commenced in 1948. Underground mining had ceased by 1954, and all production was obtained from surface operations. Innovation has been the key to Miami's success. Development of sulfide flotation and the ferric cure process for vat leaching were advanced at Miami. In situ leaching of oxide material, remaining after block caving had removed the sulfide ore, was also used.

Jan 1, 1993

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

Jan 1, 1993

By H. Langlois

"Included in the Mines Gaspe mining complex located at Murdochville, Quebec, is a copper smelter which comprises a reverberator furnace, two Peirce-Smith converters, an anode furance, and a sulphuric acid plant.The 50% smelter capacity available is supple mentedby custom concentrates for treatment. As these concentrates can be varied and extremely complex many modifications and improvements have been brought about to optimize and make the process more flexible since its inauguration in 1955.IntroductionThe copper smelter of the Mines Gaspe mining complex is located in Murdochville, a small municipality in the heart of the Gaspe peninsula, 90 km northwest of the town of Gaspe .This metallurgical installation has a daily smelting capacity of 650 tonnes of concentrate and an output of 250 tonnes of copper and 900 tonnes of sulphuric acid. Its main components are: a reverberatory furnace, two Peirce-Smith converters, one rotary anode furnace and one acid plant. The gases, which are produced by the converters, are treated at the acid plant then combined with gases from the reverberatory furnace before being vented through a 152 metre high stack .From the beginning of its operation until the present day (1991), the smelter has treated 7 million tonnes of concentrate and 100 thousand tonnes of scrap for a production of 1.8 million tonnes of copper and 2.2 million tonnes of sulphuric acid."

Jan 1, 1993

By E. G. Kelly, D. E. Walsh

The behavior of gold particles on a spiral was studied using radioactive gold tracers. The method allowed the effects of particle size, particle shape, pulp density, feed rate and splitter settings to be studied. Although the method has been successfully applied to other gravity concentrators, it was found that reliable results were more difficult to obtain on the spiral. However, these preliminary results show that, with suitable precautions, the technique is capable of providing valuable information, unobtainable from conventional studies, that could be used to better optimize spiral performance.

Jan 1, 1993

By Ian Longson, Darron W. Dixon

This paper describes the results of an investigation into the application of a statistical method for the prediction of general air-body methane concentration in longwall coal districts. An important and necessary part of the research was the acquisition of representative monitored mine environmental and coal production data. The monitored data was used to build time series models of general air body methane concentration, air velocity, barometric pressure, coal production and methane drainage variables of differing timescales according to the Box-Jenkins method of time series analysis. The univariate models were used to identify causal relationships between methane concentration and its explanatory variables. Coal production was found to be the dominant variable in the determination of the quantity of methane emitted and where appropriate, multivariate time series models were built in which expressions for methane concentration in terms of coal production were obtained. Finally, forecasts of methane concentration values were generated from the multivariate models to determine their forecasting capabilities.

Jan 1, 1993

By M. J. McPherson, P. Mousset-Jones, G. Danko, I. M. Konduri, F. Calizaya

The flow of air through a bare rock, underground airway is affected by the size and frequency of the micro- and macro-asperities of the rough surface. The focus of this study is to characterize this roughness using the coefficient of friction, the friction factor and the fractal dimension. The rough surfaces were identified by two sets of 177 and 78 transverse profiles measured by a laser profiler along the length of an underground airway at the Sunshine Mine, Idaho. Based on the measurements, the friction factor was estimated from Atkinson's equation and the coefficient of friction from the Moody diagram, and power series and variography were used for computing a fractal dimension for the airway surface. Significant conclusions are reached relating the effect of error in the measurement of cross-sectional and rubbing surface areas of a drift, on the estimated friction factor for the airway. The potential role of the fractal dimension as an alternative to the friction factor for characterizing the roughness of a mine airway is discussed.

Jan 1, 1993

By Y. Topkaya, Lloyd Petrie, Dianne Marozas, R. E. Johnson, Steven Paulson, J. H. Canterford, B. R. Palmer, R. B. Bhappu, Lorin Redden

This paper describes a fundamental approach to in situ leach mining (ISLM) lixiviant development which emphasizes leaching selectivity through detailed characterization of mineral-solution interfacial chemistry and post-leach geochemistry. This experimentation and modeling approach consists of: (1) petrographic identification of reactive ore and gangue minerals; (2) determination of leaching kinetics of individual minerals through batch leaching tests; (3) evaluation of bulk solution geochemistry through column leaching tests; and (4) ISLM simulation of element species transport and geochemistry through core leaching tests. Recent results confirm good leaching selectivity for manganese, but reveal post- leach precipitation and permeability loss during core leaching experiments.

Jan 1, 1993

By Yong-Ming Jiang

Longwall mining systems are extremely complex and expensive and require great caution for reliable and efficient operation. In many cases, ground control problems are the primary concern in longwall mining. For example, unexpected threats such as roof falls or shield damage can cause sudden interruption of coal production and safety problems, which can originate from many parameters such as shield capacity, setting load, shield type, conditions of immediate and main roofs, pillar stability, geological anomalies, etc. However, it is difficult to correlate ground-control problems at a longwall face with the related mining and geological factors. In this paper, a method of correlation analysis, called tree classification, which permits the use of a large and variable number of features, is introduced The method provides an effective approach to produce an accurate classifier or uncover a knowledge structure, which will be used to predict future consequences depending on past and current data sets. The application of the tree classification method in longwall stability analysis is discussed, and the major procedures of constructing a knowledge structure are included as an example utilizing monitored leg pressure data in a longwall face. Also, the importance and ranking of the variables related lo longwall shield stability is discussed This is a major step towards developing an artificial intelligence system for real-time interpretation of shield monitoring data and hazard prevention.

Jan 1, 1993

By T. A. P. Sankar, S. Gundiah, Pradip, R. A. Kulkarni

Two modified polyacrylamide flocculants, namely polyacrylamide with glycolic acid functionality (PAMG), and a copolymer of acrylamide with N-acryloyl-[T]- aminobutyric acid (HPVP), were synthesized, characterised and evaluated for selective separation in apatite/quartz and apatite/dolomite/calcite systems, respectively. PAMG flocculant was found selective in the separation of tricalcium phosphate (TCP) from its mixtures with quartz at acidic to neutral pH range. The selectivity deteriorated at alkaline pH. In contrast, polyacrylic acid (PAA) flocculant was found most effective only in the alkaline pH range. HPVP flocculant (even though flocculating all three minerals namely apatite, dolomite and calcite) was observed to be quite selective for the separation of calcite from apatite-dolomite minerals in single mineral tests in the presence of sodium hexametaphosphate dispersant at neutral pH. There was, however, no selectivity observed during tests on synthetic mixtures. This loss in selectivity is attributed to the phenomenon of hetero-flocculation, as reported earlier, for separation in an apatite-dolomite system with polyethylene oxide.

Jan 1, 1993

By Charles L. Kimbell

In broadest overview, a substantial part of the world's mineral industry in 1990 registered small gains over its 1989 performance, but traditional statistical measures of mineral industry performance, namely global production, trade, and consumption levels, failed to mirror these gains. This was because of the sharp downturns in the mineral industry activity in the U.S.S.R. and the former centrally planned economy nations of eastern Europe. These downturns in this one geographic area had influences on almost all traditional statistical series, in some cases simply significantly reducing growth, in other cases providing for a global downturn, in almost all cases despite considerably brighter outlooks based on activities in other world areas. To be sure, the U.S.S.R. and eastern Europe were not the only places on the planet suffering problems of one type or another in 1990 that impacted on minerals industry activity, but with the acquisition of information presently available, it seemed almost assured that the difficulties of these economies in transition will be the most difficult of problems to resolve in the long run. It certainly appeared that the legacy of the half century or more of economic central planning will be a decade or more of very difficult adjustments, and a number of these will greatly affect not only the minerals industry of these transition countries, but also the rest of the globe's mineral industry as well.

Jan 1, 1993

By Xuanxue Mo, Kaihui Yang

"The Laochang mine, located in the Sanjiang area of southwestern Yunnan Province of China, is a volcanogenic massive Pb-Zn-Cu sulfide deposit. It is hosted by a pyroclastic sequence of Early Carboniferous age, namely the lower member of the Yiliu Group, which was intruded by syenite and rhyolite porphyry. The volcanic suite is basaltic or basaltic-andesitic in composition, has alkaline affInities, and is associated with the initial rifting along the margin of the Baoshan-Shanstate continental microplate (Yang, 1989). The ores are typically Pb>Zn>Cu in composition, averaging 3.820/0 Pb, 3.51% Zn, from 0.01% to 1.85% Cu (0.107% av.), a high average content of Ag (113 glt) and recoverable amounts of Au, n, Gd, Ga, Ge, In. A massive sulfide zone, in which lenticular and stratiform orebodies are localized in a sulfide-<:arbonate-silica shale unit or sulfidic tuffIte horizon that usually overlies coarse-grained pyroclastic rocks, is underlain by a stringer sulfide zone developed in a subvolcanic intrusion or volcanic sequence. The sulfIde mineral assemblages, from the top to the bottom of the deposit, are: realgar + orpiment - galena + sphalerite + pyrite - chalcopyrite + pyrite + pyrrhotite + arsenopyrite. Oxide and sulfate minerals occur rarely in the massive ores. The massive ores are generally fme-grained and banded, but coarse-grained ore, colloform and framboidal texture, and growth-zoned pyrite and sphalerite are common. Sedimentary features such as graded bedding, brecciated and deformed ore clasts, and chaotically emplaced blocks, are well preserved, indicating that the massive orebodies have undergone slumping and displacement. Alteration is weak in the hanging-wall rocks, but strong in the stringer zone. Two distinctive alteration zones have been identifIed: the upper zone, characterized by an assemblage of quartz, carbonates, chlorite, albite, zeolite and sericite, and the lower zone containing diopside, tremolite, actinolite, gamet, clinozoisite, and epidote.The general features of the Laochang deposit are in accord with the descriptive model of VMS deposits (Lydon, 1984). However, it falls into neither the Zn-Cu group nor the Zn-Pb-Cu group chemically (Franklin et al., 1981). The depositional setting of the Laochang deposit is quite different from Cyprus or Besshi-type massive sulfide deposits. Laochang is hosted by alkaline lavas that were erupted in a submarine continental rift. Although some features of the Laochang deposit are similar to those of Kuroko-type deposits, there are remarkable differences in the alteration mineral assemblage, the proportion of ore-forming elements (lead enriched) and the amounts of oxides and sulfates. These suggest corresponding differences in ore fluid composition and the associated ore-forming environment. Thus, the authors suggest that the Laochang deposit may be a distinct type of VMS deposit."

Jan 1, 1993

By N. Rajmohan, K. T. Jacob

Presented is a thermodynamic feasibility analysis of extracting base metal chlorides from low-grade, multi-metallic oxide ores using CaC[1]2 as a chlorinating agent in the presence of SO2 and O2. The oxides react to form corresponding chlorides, while CaC[1]2 is converted to CaSO4. The Ellingham diagram is used for comparing the standard Gibbs&apos; free energy change for the sulfation-chlorination reaction of a large number of oxides. Except for alumina, silica and chromia, most of the other metal oxides will be converted to their respective chlorides. The volatile chlorides can be condensed, and the chlorides present in the condensed state can be leached. A process is proposed that uses a nontoxic chlorinating agent and gives an efficient separation of the metallic values from the gangue.

Jan 1, 1993

By Robert A. Jankowski, George E. Niewiadomski

INTRODUCTION The utilization of longwall mining technology in the United States has steadily increased since the first longwall plow system was installed in 1952. At the errs of FY 1992, some 76 coal mines were employing longwall mining systems for coal extraction, and approximately 78 shearer and 3 plow faces are operating nationwide at any given time. Although the number of operating longwalls has remained relatively constant over the last five years, longwall production levels have increased significantly during this period. During the same timeframe, the number of continuous miner sections has decreased by 13 pct., from approximately 1438 in FY 1988 to less than 1250 in FY 1992. According to Mine Safety and Health Administration (MSHA) data, the average production of longwall sections increased from approximately 2400 tons per shift (tps) in FY 1988 to more than 3100 tps FY 1992. In fact, the average production has more than tripled since 1978, when it was only 850 tps. Approximately 25 pct. of operating longwalls are capable of producing in excess of 4000 tps. Longwall production currently accounts for nearly 36 percent of U.S. underground coal production, compared to 4 pct. in 1978. These improvements are attributed to improved longwall equipment and increased availability, larger longwall panels and automation. While longwalls are highly productive and offer other advantages, operations employing this method of mining continue to experience dust compliance problems more frequently than other mining system. Although dust levels have remained relatively unchanged over the past five years, averaging 1.6 mg/m3 and 1.9 mg/m3 based on results of operator and inspector sampling, respectively, during FY 1992, approximately 25 pct. of the operating longwalls were in noncompliance during any given bimonthly sampling cycle. Although modest gains in longwall dust control technology have been made, these have been overshadowed by significant increases in production rates. As more coal is mined, more dust is generated (Figure 1). Therefore, the increase in longwall productivity has meant that far more dust is being produced which must be controlled. [ ] In view of the current compliance record, any efforts to further increase longwall output, as has been forecasted, must be accompanied by: (1) the reduction of dust levels at existing operations through the proper installation and maintenance of suitable engineering dust controls; and (2) the advancement of dust control technology for today&apos;s and future longwalls that will enable all face personnel o work in an environment that is free of excessive levels of respirable coal mine dust. MSHA&apos;S RESPIRABLE DUST CONTROL PROGRAM Since 1972, federal regulations have required each coal mine operator to continuously maintain the average concentration of respirable dust in the mine atmosphere during each shift in the active workings at or below 2.0 mg/m3 through the application of environmental dust control methods. The regulations provide for further reduction in the dust standard whenever the

Jan 1, 1993

By J. C. Kerkering, S. M. Miller, P. C. McWilliams

Researchers at the U.S. Bureau of Mines investigated the use of close-range photogrammetry and subsequent stereo digitizing to obtain data from rock fracture roughness profiles. The photogrammetric process yielded results that were acceptable but somewhat inferior to those obtained by a mechanical profilometer. On the basis of this study, further pursuit of photogrammetry as a data collection method in mining is proposed. Fractal geometry was investigated as a means of measuring the roughness of rock fracture profiles. Four fractal algorithms were used: divider method, modified divider method, box method, and spectral method. A comparison of the methods gave ambiguous results. Brown&apos;s modified divider method provided the best means of obtaining the fractal dimension. Shear strength estimates were obtained using the parameters of the modified divider method and Myers&apos; Z, measure. Because of differences in results when comparing the different ways of obtaining the fractal dimension, future users of fractals in studies of rock fractures are advised to cross-check their results carefully.

Jan 1, 1993

By G. P. Larson

We define a specialized mineral market as follows: Specialized markets occur where a low volume of a given mineral is used to convey a large benefit to a specific product. Sales of these value added minerals to small markets generally require a different marketing approach than bulk or commodity markets for traditional minerals. Specialized markets are differentiated by several characteristics as listed in Table 1. [Table 1-Characteristics of Specialized Markets 1. Long-term sample evaluation 2. Small volume markets 3. Highly controlled properties 4. Customized specifications 5. Technical selling effort] Many of the specialized markets require long term evaluations before sales can be made. This can vary greatly depending on the type of industry and ultimate end-use of the final product. In some cases, complete evaluations may take from several weeks to years. Length of sample evaluation has no relationship to size of market. That is, the total market may be quite small or fairly large (e.g. from a few hundred pounds to a few million pounds). Uniformity of value added minerals from lot to lot is an extremely important characteristic for these markets. Variations that are of no consequence to one industry (or application) cannot be tolerated in another. This has led to customized specifications and test procedures for specific markets and applications. Because of the greater uniformity and customized specifications, the need for greater technical selling effort is required. This is particularly true in the initial developmental phases and contacts as well as in maintaining the account after the initial order. A small entrepreneurial organization can be used to advantage to achieve goals of sales and profit to these types of markets. The large commodity or bulk producer&apos;s sales and marketing organizations are not ordinarily equipped to give the necessary service or technical assistance to the specialized markets. The organization usually required for this type of marketing is given in Table 2. [Table 2- Specialized Marketing Organization Requirements 1. Flexible approach to technical and marketing problems 2. Versatile background in a broad range of industries 3. Laboratory backup when necessary 4. Marketing expertise 5. Technically competent management] In developing specialized markets, a flexible approach to both technical and marketing problems is necessary in understanding the needs and requirements of the industry and the applications involved. This flexibility is also helpful in adapting existing products to new market areas. Having knowledge of a range of industries as broad as possible can be very helpful. It often happens that a product developed in one area or industry can lead to its application in a different area with excellent results. Laboratory assistance for a wide range of tests and test procedures is an important asset in any development program for specialized markets. Laboratories can also assist in competitive evaluations. Marketing expertise and a management with some degree of technical understanding are necessary in to assign proper cost/ benefit ratios to a proposed new product or application of existing products. This is important since the total market volume can be small and the time required for initial contact to first sale can often be lengthy. Also, management&apos;s technical and marketing expertise is required in evaluating all possible options that can be considered viable when an application is presented. Once an organization of this type is in place, it is then possible to determine the needs of the marketplace, including when various minerals might have to be customized to fill those needs. There are generally two methods of marketing, as noted in Table 3. [Table 3- Method of Marketing 1. Product Driven -Develop a product with certain desirable properties and try to find a market to utilize those properties. 2. Market Driven -Know market well enough that when a need arises, it is possible to identify a product that will satisfy the need.] These approaches are not mutually exclusive and can be used at the same time. Marketing a value added mineral generally requires the "market driven" approach. Hence, the importance of a knowledgeable management as outlined above. To determine where needs and opportunities exist, several approaches can be used. Table4 lists some of the methods employed.

Jan 1, 1993

By Shinnosuke U, Zhang HE, Qian X

In this paper it is recommended that pyrite is to be separated as froth from arsenopyrite by selective depression of arsenopyrite in strong alkaline medium (1), by oxidation with chlorinated lime (II), by co-use of chlorinated lime with sodium carbonate (III), and by modification with `L&apos; reagent (know-how) without filtering to get rid of residual copper ions after Cu activation (IV). The corresponding principles of these (new) effective methods were briefly discussed. The artificial mixture (1:1) of pyrite and arsenopyrite could be successfully separated, and concentrates of pyrite containing 2.12 per cent (I), 7.00 per cent (II), 2.60 per cent (III) and 3.16 per cent (IV) arsenic with recovery of 90.9 per cent (1), 89.5 per cent (II), 89.0 per cent (III) and 81.4 per cent (IV) respectively can be obtained by only one flotation. Some effective precautions were suggested to improve the separation of arsenopyrite and pyrite (or other sulphides) by flotation.

Jan 1, 1993

By R. L. Grayson, R. C. Althouse, L. A. Layne, M. J. Klishis

Introduction The researchers combine a detailed microanalysis of accident data and roof bolting field observations to develop new and meaningful risk indices for the roof bolter job classification. These indices incorporate information that has not been used in traditional measures of risk. The significant parameters that these new indices incorporate are the total number of accidents occurring within a particular work activity, the amount of time a miner usually spends performing the work activities in which the accidents occurred, and the average severity of the injuries incurred in the accidents. Being specific for a particular job, they aid comparisons among work tasks performed in the bolting cycle. The indices&apos; specificity provide anew basis for reexamination of training and operational development of miners for specific jobs. Microanalysis The roof bolting microanalysis was performed by examining the empirical data and written summaries for all bolting accidents reported in the West Virginia Safety Information System (WVSIS) for the years 1983-1987. The WVSIS database contained more than 15,000 accident narratives. The microanalysis of bolting accidents covered 2083 reported personal injuries (RPIs). This large number of injuries, performed at an industry-wide level, provides a broad picture of the nature of roof bolting injuries. The analysis focuses on the activities associated with the cyclical routines of the roof bolting process. For this analysis, roof bolting tasks were categorized into work routines based on the sequence of work activities performed by roof bolters. Four work routines were identified within the overall bolting cycle. And these routines comprised 70% (1217) of all roof bolting accidents at the face. The routines include: face area preparation; tramming, positioning and setting the automated temporary roof support (ATRS); drilling holes; and installing bolts. The types of injuries most common within a routine and the average severity of those RPIs were also identified. Table 1 gives a breakdown of the percentage of accidents by the bolting routines. Another 19% of the roof bolting RPIs cannot be classified in a particular routine because of the vagueness of the narratives. However, it was clear that these accidents involved drilling holes or installing bolts. These results are consistent with previous research. It identified drilling and bolting, followed by tramming, as the most frequent sources of injuries (Helander, Krohn & Curtain, 1983). [Table 1-Accidents by bolting routines at the face Accidents Bolting routine Percent Number Face area preparation 18.2 222 Tram, position, ATRS 23.3 284 Drilling holes 34.2 416 Installing bolts 24.2 295 Total 1217] Roof bolter operators were involved in 82% (1706) of the 2083 accidents. The RPIs incurred by classified roof bolters occurred largely on working sections (95%) and at the face area on the working section (75%). Approximately 23% of the 1706 roof bolter RPIs occurred during tasks which were not part of the bolting cycle. These included belt utility work, providing supplies and section moves. Miners in other job classifications who were involved in roof bolting accidents accounted for the remaining 18% of the 2083 roof bolting RPIs. Among this group were section utility men and general laborers (41%), section machine operators (38%), and electricians and mechanics (11%). Another 9% were performing on-section supervisory functions. Face area preparation Face area preparation work led to a substantial number of injuries, although it accounted for the lowest number of injuries of the four routines. Several tasks were associated with the injuries incurred in this routine. Table 2 shows the percentage of injuries associated with face area preparation. [Table 2- Accidents occurring during face area preparation 26.6% Setting temporary and supplemental supports 24.8% Scaling roof and ribs 21.2% Installing ventilation devices 15.8% Handling supplies and materials 11.6% Miscellaneous tasks] The injuries commonly involved overexertion and &apos;struck by&apos; falling coal and rock. Handling materials often led to overexertion injuries. Nearly 66% of the scaling-related injuries and about 33% of the injuries incurred during the setting of temporary supports involved &apos;strikes&apos; by equipment, materials, top and rib. Tramming, positioning and setting the ATRS During tramming operations, workers were often struck by components of the machine and hit by loose roof and rib. Poor footing also led to slips, sometimes against the moving machinery. Table 3 gives a breakdown of accidents in this routine. Control levers were a significant problem in this routine.

Jan 1, 1993

By William J. Maloney

OVERVIEW Blind drilled shafts are becoming increasingly popular, particularly in the Appalachian coalfields, as a secondary and sometimes primary means of ventilation. The reasons for the rebirth of the blind shaft are mainly due to it&apos;s economics, timeliness, safety characteristics and inherent independence from underground operations. Historically blind drilling has been used to a limited degree in hard rock and tunneling ventilation. With the recent advances in cutters, bit design, equipment and lining systems, particularly in the Appalachian coalfields, blind drilled shafts are once again the shaft construction technique of choice to the educated ventilation specialist. [ ] HISTORY OF BLIND DRILLING Blind drilling has it&apos;s roots in numerous places, including the Nevada Test site, the coalfields of Appalachia and Germany along with the uranium mines of Arizona. Early technology was developed using a variety of equipment mainly of oilfield origin with oilfield drillers entering the industry on a project basis only to "pull in their horns" and return to the oil patch. Shafts completed by oilfield equipment range from the shallow 5&apos; and 6&apos; lined shafts on the Nevada test site to two (2) 10&apos; lined, 730&apos; deep shafts in Alberta for heavy oil development. Several of these shafts were completed to extraordinary depths with the record being a 90" diameter, 6,150&apos; deep shaft being completed in the Aleutian Islands of Alaska in 1969. EQUIPMENT Oilfield equipment is still being used on the Nevada test site but elsewhere specialized blind hole equipment has been developed to enhance the efficiency of particular applications. Where steel liners are required the A-leg derrick with a simple hoist and rotary system is used. The largest user of this type equipment in the world is the former USSR where blind drilling for mine ventilation is the norm. On applications where trip time and cutter changeout are a priority, top drive hydraulic rigs have been developed which are highly automated, efficient and mobile. These rigs are particularly well suited for tunneling applications where shafts are shallow and sites are usually tight.

Jan 1, 1993

By Floyd C. Bossard

A method of hand-calculating the air-cooling requirements of a conceptualized underground mining operation is presented for the reader&apos;s orientation. Separate air heat load calculations were conducted for adiabatic compression, electromechanical equipment, wallrock, broken rock, groundwater, and blasting operations. The air heat sources were calculated for four mining levels under conditions representative of a typical planned mining operation at depth. The total heat gain on each level will approximate the level air-cooling requirements. The results of these hand calculations can be further modified by the use of mine ventilation computer software that refine the heat-source calculations, predict underground ambient air temperatures, and establish the air-cooling requirements of a mine. INTRODUCTION The principal mining method employed is a modified vertical crater retreat (VCR) blasthole operation. Typical scopes range from 20 to 40 feet wide, 75 to 100 feet long, and 100 feet in height. Six and one-half inch blastholes are drilled with an "in the hole" hammer drill. Eight-foot high horizontal rounds are blasted down. Mucking of ore from the undercut to orepass is done with LHD equipment operated from a remote control station. Backfill includes hydraulically placed tailings with cement, and waste rock when available. Stope access is from ramp sub-levels on 50-foot vertical intervals. Crosscuts are ramped down to the first stope cut 25 feet below the sub-level elevation. Then the crosscuts are raised by taking down the back when each stope cut is completed, until an elevation 25 feet above the sub-level is reached. The crosscuts are filled with tailings, and/or waste rock. See Figure 1. Typically, a two-pronged approach to defining the air-cooling requirements is conducted. First, the principal sources that make up the air heat load are individually hand-calculated. Second, projections of mine heat load are calculated by utilizing computer modeling techniques. This paper discusses the first method (hand calculations) of determining the individual components of heat flow into the mine. CALCULATED AIR HEAT LOADS Adiabatic Compression The plans for the mine include delivery of 300,000 cfm of air to ventilate the lower level operation. This is equivalent to approximately 200 cfm/ton of rock produced (300,000 cfm/1500 tpd of ore and waste). As air flows down a shaft, with no heat interchange between the shaft and air and no evaporation of moisture taking place, the air is heated in the same way as if it were compressed in a compressor. Dry air increases in temperature about 5.4°F per 1000 feet. One BTU is added to each pound of air for every 778 feet of decrease in elevation, or is subtracted for the same elevation increase. For dry air, the dry-bulb temperature change is 1/(0.24 x 778) = 0.00535°F/ft., or 1°F/187 ft. elevation. Auto-compression may be masked by the presence of other heating or cooling sources, such as shaft wallrock, groundwater, air and water lines, electrical facilities, etc. The major factors influencing the temperature of the air delivered underground by a shaft are (1) the night time cool air temperature&apos;s effect on the rock or lining of the shaft, (2) temperature gradient of ground rock related to depth, and (3) evaporation of moisture within the shaft which increases the latent temperature and decreases the sensible temperature. [Calculation For Adiabatic Heat of Compression] a. Assumptions: 1. Three hundred thousand cfm of fresh air at 3000 level has increased in temperature during the summertime to the point where it has little available cooling power. Air-cooling will be required on 3000 Level and below. 2. Elevation of 3000 Level is approximately +1000 ft. above sea level.

Jan 1, 1993

By E. Oshima

The only one reverberatory furnace smelting in Japan has been alive in Onahama Smelter, Onahama Smelting and Refining Company. The fuel for the reverberatory furnaces had been changed from bunker C-oil to mainly coal and partially waste tires to reduce the energy cost. The steam, recovered by the waste heat boilers of the reverberatory furnaces, has been utilized fully for the power generation to save the purchase power. Copper concentrate has been smelted also by injection through tuyeres in the P.S. converters since 1983. The amount of concentrate, treated in the converters, was increased by the construction of small oxygen plant in December, 1990. In the paper, the current practice and improvements in the reverberatory and converter operations are described.

Jan 1, 1993