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By Kremena Dedelyanova, Tsvetan Damyanov, Nikolay Hristov, Evgeni Kostadinov, Lubomir Kuzev

"Spherical grinding media (balls) are the most widely used media for comminution of ores by grinding. One way to optimize the grinding process in a tumbling mill is associated with the development and testing of a new media shape. By modelling simulation study of the behaviour of two types of grinding media (balls and tetraballpebs) in the grinding process was conducted to determine the possible advantages of the new media shape. Tetraballpebs are in the shape of tetrahedrons with spherical convex walls. D305 mm and L305 mm Bond ball mill sizes were used in the experiment. They are modelled milling tetraballpebs bodies corresponding by weight to five ball sizes, namely ¾, 7/8, 1, 1¼ and 1½ inches which form the grinding media used in a Bond ball mill. They are cast steel grinding bodies. Studies carried out with both types of grinding media include equal number of milling bodies of the same weight and volume. Having thus ensured a uniform mass of the grinding media and the same number of milling bodies, all other conditions and process parameters being identical, conclusions can be drawn about the influence of the grinding media shape on the grinding process. Compared to steel balls, the tetraballpebs used in laboratory tests carried out in closed cycle by Bond ball mill and the study of the rate of grinding gave five to six per cent higher yield of a control size of minus 0.1 mm. Measurements of the Bond ball mill power consumption demonstrated that tetraballpebs grinding media achieved energy saving of 4.7 per cent for the production of one gram of the control size of minus 0.1 mm."

Jan 1, 2014

By Charles P. Williams

The lead-bearing area of Missouri has been subdivided, geographically, into the Southeastern, the Central, and the Southwestern

By A. C. Smith

The U.S. Bureau of Mines conducted a study to evaluate the effect of ventilation on the water spray patterns of automatic sprinkler heads. Experiments were performed in a rectangular tunnel with pendent, upright, pendent sidewall, and horizontal sidewall sprinkler heads at air velocities of 0, 150, 300, 500, and 800 ft/min. As the air velocities were increased, there were significant shifts in the total coverage areas, density distribution patterns, and the maximum coverage densities for all types of heads. The pendent and upright heads exhibited upstream shifts in total coverage in the direction of the air-flow ranging from 4 to 6 ft, while the downstream coverage distances were extended up to 14 ft at the 800 ft/min airflow. The shift in upstream coverage distance for the sidewall heads ranged from 6 to 12 ft, while the downstream coverage was extended up to 22 ft at 800 ft/min. The results showed that airflow can have a significant effect on the coverage characteristics of automatic sprinkler heads and needs to be considered in the design of effective sprinkler fire suppression systems for ventilated areas.

Jan 1, 2010

By Arthur L. Walker

In June, 1893, while I was with the Old Dominion Copper Co. at Globe, Ariz., I received an offer to take charge, as manager, of the Baltimore Electric? Refining Co. which had recently built a plant at Baltimore, Md., having a capacity of about 1200 tons of electrolytic copper per month, using the Hayden series system. This was a very large output, for in those days the plants were small and the art was in its infancy. At the beginning of 1893 there were 11 electrolytic refineries in the United States, most of them using the multiple system. The Hayden series system was used at Baltimore and Bridgeport and the Smith series system was in operation at the Electrolytic Copper Co. at Ansonia, COM., and at the Pennsylvania Salt Co. at Philadelphia. In the latter system the plates were placed in a horizontal position, but the operating difficulties were so numerous it was soon discarded. The total production of these 11 plants in 1892 was very small; I have not been able to discover reliable figures of production. There were about thirty electrolytic copper refineries in the world that year, according to Titus Ulke, and the entire production of these

Jan 1, 1932

By Neha Shukla, Nikhil Dhawan

The current study examines the extraction of rare-earth elements from trichromatic phosphor of discarded tube lights via baking processes. The Y2O3: Eu3+ (YOX) and BaMgAl10O17: Eu2+ (BAM) phases completely decomposed during acid and alkali baking with > 90% Eu and Y extraction. Prior, selective acid leaching of the YOX phase enhanced the acid baking process at a lower temperature (< 230 ℃). Feed acid baking showed a maximum 87% Tb, 36% La, and 17% Ce extraction at 500 °C, 0.5 h, 1 ml/g H2SO4, and alkali baking yielded 16.1% Tb, 86% La, and 73% Ce at 300 ℃, 0.5 h, 50 wt.% NaOH. The acid baking of the acid-leached phosphor yielded 72% Tb, 95% La, 63% Ce, ~ 95% Y, and Eu extraction at 300 °C, 0.5 h, 1.05 mL/g H2SO4. Lower Tb extraction during the acid baking of acid-leached phosphor can be attributed to the formation of Ba3Tb( PO4)3 and BaTbO3 phases. Insoluble Si3Tb5 phase formation and partial decomposition of CeMgAl11O19: Tb3+ (CMAT) phase in the alkali baking route leads to lower Tb extraction. The acid baking process is suitable for extracting Tb, Eu, and Y, whereas the alkali baking process is suitable for La and Ce. The thermal analysis of phosphor and flux interaction depicts that alkali baking is less energy-intensive than acid baking. In comparison, the one-step acid baking process and the two-step alkali baking process yielded desirable extraction results (87% Tb, total RE (rare earth) 78.5% and 50.4% Tb, total RE 59.3%). The product yield in the acid and alkali baking route of feed, acid-leach phosphor is 36%, 29%, and 39%, 28.2%, respectively, with ~ 98% mixed rare-earth oxide purity.

May 5, 2022

By O. C. Ralston, William C. Hunter

DUring experimental study on the separation of a zinc concentrate from the heavy sulfide ores of Jerome, Ariz., a great deal of attention has been given to proper activation of the sphalerite content of the ore. The sphalerite at Jerome contains enough iron to class it as a marmatite, as picked specimens have only about 60 per cent. zinc, whereas pure zinc sulfide contains 67 per cent. zinc. This marmatite in large pieces is black in color, but in dry powdered state is brown. A flotation froth made from it has a silvery black luster in comparison with a brassy luster caused by chalcopyrite which occurs in the same ores. The sample used in most of the experiments described below consisted of about 3000 Ib. of diamond-drill core from the zinc-bearing orebodies of the United Verde mine, and had the following analysis: Per Cent. Per Cent. Oz. peR Ton Cu............... 0.695 Al2O3............ 0.83 Ag................. 1.30 Zn............... 6.17 S................ 43.35 Au................ 0.02 Fe...............36.23 CaO............. 0.46 SiO2............. 9.27 Mineralogically, this analysis is calculated to the following, which is in agreement with the microscopic examination: Per Cent. Per Cent. Chalcopyrite................... 2.0 Schist........................... 3.2 Marmatite..................... 10.32 Quartz........................... 8.2 Pyrite........................ 73.0 Calcite.......................... 0.9 All samples were ground in a laboratory ball mill in order to leave about 3 per cent. on a 200-mesh sieve, fine grinding being essential to liberate the minerals in the ore. Most of the Jerome ores are so intimately crystallized that reasonably complete unlocking by grinding does not take place until 800 mesh is approached. Therefore, the 200-mesh grind adopted contains a considerable amount of locked mineral grains, but not so much as to render impossible the preparation of commercial grades of concentrate.

Jan 1, 1930

By C. E. Gregory

In terms of a Research Proposals supported by a Grant Agreement dated 30 June 1973, the author was commissioned to prepare an annotated bibliography of mine fires. The main objective was to conduct a comprehensive program of library research and search of official files to prepare a list of articles dealing with actual occurrences of mine fires. This bibliography as now completed includes an annotated reference from each article or report cited, underlining the cause, extent, methods used, procedural errors made, results achieved, ventilation effects, and lessons learned from the operation of subduing the particular fire.

Jan 1, 1974

By J. W. Freeman, E. E. Reynolds, A. E. White

Fram a study of 63 systematic alloy modifications it was found that molybdenum, tungsten, and columbium, added individually or simultaneously, and increases in chromium cause major improvements in 1200°F rupture strengths of Cr-Ni-Co-Fe base alloys. Rupture strengths were a function of the effect of composition modifications on both the inherent creep resistance and the amount of deformation the alloy would tolerate before fracture. THIS paper describes the results of an investigation of a series of alloys with systematic variations of the chemical composition of the following basic alloy: C, 0.15: Mn, 1.7; Si, 0.5; Cr, 20.0; NI, 20.0: Co, 20.0: Mo, 3.0; W, 2.0; Cb, 1.0; N, 0.12; Fe, 32.0 pct. The 62 modifications of this alloy were produced under conditions which minimized all factors influencing properties at high temperatures except composition. Melting, fabrication, and heat treatment were carefully maintained constant. Stress-rupture properties at 1200°F were used as the primary criteria of evaluation of the alloy. The objective of the study was to obtain data for determining the fundamental role of the influence of alloying elements on properties of heat-resistant alloys at high temperatures. In addition the results should be useful in determining optimum chemical compositions, the sensitivity of properties to variations in composition, and the degree to which alloy content could be reduced while retaining worthwhile properties. It is difficult or impossible to develop correlations between properties at high temperatures and systematic variations in chemical composition from published data for wrought heat-resistant alloys developed for gas turbines.&apos; &apos; The main reason for this is the extreme dependence of the properties on conditions of treatment of the alloys." In most cases variation in final treatments between alloys so influences the properties that the influence of chemical composition is obscured. In addition it is recognized Table I. Basic Alloy and Some Modifications Used Basic AllOy, Variations in Element Pct Composition, Pct C 0.15 0.08. 0.40. 0.60 Mn 1.1 0.03,0.25.0.50,1.0,2.5 S1 0.50 1.2, 1.6 Cr 20.0 10, 30 Ni 20.0 0, 10,30 Co 20.0 0. 10, 30 MO 3.0 0. 1.2.3, 5, 7 W 2.0 0, 1, 5, 1 Cb 1.0 0.2,4,6 N 0.12 0.004, 0.08, 0.18 Fe 32.0 that variations exist between heats of the same alloy which are related to melting practice and that there is a strong possibility that conditions of hot working influence response to final treatments. The development of heat-resistant alloys has been based on the gradual accumulation of data roughly related to composition from extensive testing programs. There is every reason to believe that in most cases the optimum compositions have been achieved by this procedure in the alloys commercially available. There are, however, very little data showing the influence of systematic variations of composition free from the influence of other factors, particularly for alloys of the type investigated. Several investigators of cast alloys have demonstrated compositional effects, notably Grant,1-6 Epremian,t Guy,8 and Harder and Gow.9 Sykes10 eviewed the work on the wrought alloy Rex 78 and the systematic variations of carbon, copper, molybdenum, and cobalt leading to the development of the stronger Rex 337A alloy. From the papers by Wilson11 and Henry12 it is possible to deduce the beneficial effect of substituting cobalt for iron in 0.45 pct C-20 pct Cr-20 pct Ni-4 pct Mo-4 pct W- 4 pct Cb alloys. Wilson mentioned but did not present the extensive compositional studies involved in developing these alloys. Binder" showed optimum properties for 3, 2, and I pct, respectively, for molybdenum, tungsten, and columbium in 20 pct Cr-20 pct Ni-20 pct Co-30 pct Fe alloys for limited systematic variations of these

Jan 1, 1953

By Satoshi Honda, Mitsuru Shimizu, Tatsuji Nagao, Motoaki Kurisu, Hirokazu Izumi, Toshinori Takahashi, Hiroyuki Itoi

"1. INTRODUCTION Grade separation of roads and railway tracks is an effective method for solving traffic-related issues in cities. However, in urban areas where the existing traffic infrastructure is already overcrowded, construction of new roads often faces problems especially in the proximity of existing structures. This paper summarizes the construction plan and results of new road tunnel construction in the proximity of the piers of elevated tracks on the Tokaido Shinkansen railway.2. SUMMARY OF CONSTRUCTION AND ISSUES Auxiliary Route No. 26 as specified in the Tokyo urban plan is a project by the Tokyo Metropolitan Government, covering the construction of a new road with a total length of 22.3 km currently on the way to completion. In Shinagawa Ward, Tokyo, Auxiliary Route No. 26 is designed to cross the railway tracks of the Yokosuka Line and the Tokaido Shinkansen Line, so for this project, the East Japan Railway Company (JR East) was commissioned to design and build a road crossing under the railway tracks (Sumiyoshi Underpass). Figures 1 through 3 show a sketch of the Sumiyoshi Underpass. The Sumiyoshi Underpass was planned as a two-lane tunnel for vehicles, which was to cross under the railway tracks of the Yokosuka Line and the Tokaido Shinkansen Line at an angle of 70 degrees. The structure concerned is a single-story single-span box culvert structure with the following dimensions: 11.5 m wide, 7.7 m high and 31.6 m long. The Tokaido Shinkansen Line tracks were constructed on a girder-type elevated bridge built in the 1960s, and the foundation piles of the piers (P7 and P8) in the proximity of this crossing box culvert of RC structure were constructed using an earth drill method. The crossing structure for this project is located extremely close to the piers of the Tokaido Shinkansen railway tracks, with a minimal separation of 0.4 m, and the most significant issue of this project was construction in a manner that would not obstruct the normal operations of the Tokaido Shinkansen (Figure 4). Initially, we considered installing a temporary earth retaining structure in the proximity of the piers, cutting the ground and building an RC box, but concluded that it would be very difficult to apply this method because of probable excessive displacements in the Shinkansen railway piers. That is why we decided on the use of a non-open cut method in order to minimize the impact on the piers."

Jan 1, 2016

By W. E. Anderson, T. M. Brady

As part of a larger study to identify factors influencing the practical operating life of wire rope used on large draglines in surface coal mining, field trips to operating surface coal mines were made during late 1975. Wire rope performance information was obtained from personnel at 13 mine sites west of the Appalachians. Seventeen pieces of equipment, utilizing bucket sizes ranging from 30 to 220 cu yd (23 to 168 m), were documented. Draglines at large US surface coal mines can be represented as 120 machines having 60-cu yd (46-m3) buckets, moving 1.8 x 10[9] cu yd (1.4 X 10[9] m3) of overburden during 6650 hours of productive availability per year. Based on the limited field study, the calculated downtime resulting from rope-related causes averages 105.6 hours per year for each typical 60-cu yd (43-m3) dragline. The total loss in production from this source alone is thereby estimated at about 25 X 10[6] cu yd (19 X 10[6] m3) annually.

Jan 11, 1979

By E. F. Loiseau

In a paper on the manufacture of artificial fuel, read at the Philadelphia meeting of February, 1878, I enumerated the difficulties which I had to overcome before succeeding in the mixing of coal-dust and clay, the compressing of the same mixture, and the water-proofing of the lumps. The drying of the lumps, after leaving the press, was the remaining difficulty, and it mas expected that a plan devised by Dr. Charles N. Cresson, of Philadelphia, would enable us to dry the fuel as rapidly as it was moulded, and that a continuous production could in that way be obtained. The company was reorganized. The works were purchased by the new company at an assignee&apos;s sale, and the oven was modified, according to Dr. Cresson&apos;s plan. Anticipating a possible failure, I had prepared a plan by which I expected to be able to demonstrate that anthracite coal dust nixed with pitch, could be manufactured with our present machinery slightly modified ; so that after all, if we were compelled to give up the attempt to make fuel for domestic use, there was a possibility of succeeding in the manufacture of a good steam-fuel. The plan suggested by Dr. Cresson for drying the pressed lumps of coal-dust cemented with clay, did not work as well as we expected. It enabled us to dry more fuel than we did before, but it could not be made to dry more than one half of the lumps produced by the press. The plan mas abandoned, and I was authorized to experiment with coal-dust and coal-tar pitch. The cement which is used in Europe to conglomerate coal-dust is usually dry pitch, which is prepared by separating from the tar, at a temperature of 572&apos; Fahrenheit, the volatile matters which it contains. Some manufacturers, however, employ crude tar, others, a rich tar, which has been cleared of 25 per cent. of its volatile substances, by heating it to 392&apos; Fahrenheit. But with common tar very weak fuels are obtained, which do riot burn well, and give out a strong smell and a great deal of smoke ; it is also necessary to subject them to a baking process, in order to solidify them, and to eliminate the more volatile of the materials contained. This operation of course requires a special plant, the cost of which increases sensibly the price of manufacture, without counting the products

Jan 1, 1880

By L. A. Neumeier

A refinement of the tungsten-cobalt phase diagram is presented for compositions to 85 at. pct cobalt and for temperatures to the solidus. Preparation of alloys by powder metallurgy methods is described, and the results of studies made by a combination of techniques involving microscopy, X-ray diffraction, thermal analysis, and dilatometry are discussed. The existence of u-W6Co7 is confirmed (rhombohedral D85 structure). The W6Co7 phase forms periectically at 1,685&apos;115° C by reaction of liquid with primary W; the peritectic melt is at 71 at. pct cobalt. The homogeneity range of u-W6Co7 extends to both sides of the ideal stoichiometric ratio, through a compositional range of about 5 at. pct. The compositional limits for u-W6Co7 vary slightly below 1,471°±5° C, the temperature of the eutectic reaction forming u-W6Co7 plus a. The eutectic composition is at 79 at, pct cobalt. The existence of WCo3 is confirmed (hexagonal D019 structure). WCo3 forms peritectoidally at 1,093°±5° C by reaction between a and u-W6Co7 and exists over a narrow homogeneity range of 0.5 to 0.6 at. pct. The solid solubility of cobalt in tungsten is low (maximum 0,9 at, pct cobalt) and results in contraction of the tungsten lattice. Tungsten is extensively soluble in a-Co at higher temperatures; the a-Co lattice is expanded by the presence of tungsten in solid solution. There was no indication of the formation of a o-phase nor of any intermediate phases other than WeCo7 and WCo3. The WeCo7 and WCo3 phases are compared with similar phases in other systems.

Jan 1, 1967

By Robert W. Bartlett, Keith A. Prisbrey

"Natural biooxidation of shallow, refractory ore heaps, relying on gaseous oxygen diffusion from the heap surface is attractive because of its simplicity and potential low cost, especially without using a lined pad, which is possible if percolation of leachate through the heap is avoided. In this case, the excess oxidation enthalpy must be removed by water evaporation from within the wet ore heap, and the rising water vapor flux within it will impede the downward oxygen flux. An analysis of this process shows that for typical values of pyrite grades and limited temperature increases in the ore heap, the oxidation rate and penetration into the heap is reduced to values between about 30 percent and 50 percent of those obtained when leachate percolation occurs and water evaporation is not a significant mechanism of heat removal.IntroductionHeap biooxidation of refractory gold ores, needed as a pretreatment step to liberate the gold for subsequent extraction by leaching, requires large amounts of oxygen for the normal amounts of sulfide minerals that are typically present, one to two percent pyrite and/or arsenopyrite.The importance and potential limitations of oxygen transport into commercial ore heaps, which are much larger than encountered with column tests of ore and with pilot heaps, has been recognized111 • Commercial ore heaps using radial stacking conveyors are typically about nine meters high and may be 100 meters or more in lateral dimension. Three oxygen transport mechanisms have been identified and analyzed: ( 1) natural convection, which arises from an increase in buoyancy caused by the temperature rise and gas composition changes associated with the exothermic oxidation of the sulfide minerals within the ore heap, (2) forced air ventilation, using low pressure fans and air distribution pipes within the heap, and (3) diffusion of oxygen in the gas-filled void spaces within the ore heap driven by the oxygen concentration gradient from the surface of the heap to the mineral oxidation loci within it."

Jan 1, 1997

By S. Jeon, N. Hiroyoshi, N. Kitajima, S. Tanaka, M. Ito, R. Pongsumrankul, C. B. Tabelin

"Gold (Au) extraction is inefficient when crushed waste mobile phones (WMPs) are directly leached by ammonium thiosulfate likely because of the cementation/precipitation of dissolved Au in the presence of base metals like copper (Cu), iron (Fe) and/or aluminum (Al). This study applied physical separation techniques to remove base metals as well as plastics and concentrate the Au containing fractions from WMPs prior to leaching. Coarse fraction of crushed WMP (+2-5.6 mm) was pretreated by jig and magnetic separation to remove plastics and base metals like Fe, Cu, and Al, while the fine fraction (-2 mm) was pretreated by flotation, and float and sink separation. The leaching results showed that Au extraction increased in both of the size fractions. In the coarse fraction, Au extraction increased by as much as 2-fold after pretreated (from 4 to 7%). Improvement of Au extraction in the fine fractions after pretreatments were even more dramatic. Au extraction increased from 4 to 61%. These results indicate that physical pretreatment to remove co-existing common metals and plastics is effective to improve Au extraction from WMPs. INTRODUCTION Global sales of mobile phones are expected to increase steadily in the foreseeable future because a lot of people buy newer phones with upgraded and more advanced functions almost every two years, which leads to the higher generation of waste mobile phones (WMPs). WMPs contain a wide variety of materials including plastics, glasses, base metals and precious metals such as gold (Au). Among all the valuable materials in WMPs, Au accounts for over 80% of its total metal price (Park et al., 2017), so recovery of Au from WMPs is needed."

Jan 1, 2018

By José Ricardo Lopes, Alaor Coelho, Marco Gomes, Alfredo Dacio Moraes Filho, Akira Koshima

"Very low strength alluvial peat soil was found under two old drainage galleries with just 4.3m cover above a twin tunnel due to be excavated by a 10.6m diameter EPB Shield for São Paulo Metro’s Line 5. Alignment was required to strengthen soil and drive the TBM through safely without damaging the galleries. The design included six levels of self-drilling forepoling with lengths up to 15m, using 100 mm-diameter SCH160 metal tubing and injection bulbs with estimated 0.20m diameter. Two 6m-diameter circular shafts were excavated along the shield-tunnel axis, one each side of the galleries, and forepoling fanned out under gallery foundations. Given the peculiar soil and confined working space, more data were needed for correct drilling geometry. Ground mass treatment was successfully designed and executed and galleries were fitted with instruments to enable the shield’s advance. DESCRIPTION São Paulo Metro’s Line 5 (aka Lilas or purple line) is adding 11 stations along an 11.5km extension and three EPB TBMs are being used simultaneously (three EPBs on the same job for the first time in Brazil) with one of them excavating a large cross- section twin-track tunnel. Currently the largest work-in-progress for São Paulo Metro, this project has been divided into seven lots and 9km are already operational. This study refers to Lot 6 where the 10.6m - diameter EPB shield excavation is only 4.3m below two galleries running through a 10m layer of very soft peat soil in an old alluvial riverside region ringed by major roads. After recent filling, this area (Parque das Bicicletas) was used for recreation, although temporarily it has become a construction site. This paper will describe the original basic design; precautions in relation to the structural condition of the galleries after inspection; the problem of absence of data for their foundations; subsoil bore tests; evolution of the engineering design for the galleries and conditioning and reinforcing the soft soil below the galleries to ensure safe advance of the EPB shield, and finally the successful outcome of the solution used."

Jan 1, 2016

By Kun Li, Shuaishua Yue, Huigui Li, Yuan Ruifu, Yongxiang Xu, Huamin Li

"Coal seams in Tashan Mine of Datong Coal Group in China average 49.2 ft (15m) thick and have been mined by the top coal caving longwall mining method of large mining height. Mining height was 12.5 ft (3.8m) and the top coal caving height was 36.7 ft (11.2m). The gateroad pillar between panels was 124.6 ft (38m). During retreat mining, serious bumps occurred in the gateroads on both sides of the pillar affecting safety production. Therefore, pillarless mining was experimented.Using numerical modeling and comparative study of cases of similar mining condition, it was decided to employ a 19.7 ft (6m) wide pillar, rather than the previous 124.6 ft (38m) wide pillar. Support system for the gateroads was designed and implemented. During gateroad development, pillar failure conditions and entry deformation were monitored. Hydraulic fracturing method was employed to cut off the K3 sandstone along the entry rib so as to reduce the abutment pressure induced during retreat mining. Support reinforcement method combining grouting and advanced reinforcement methods was proposed to insure stable gateroad ahead of mining. Methane drainage and nitrogen injection were implemented to eliminate hazards associated with mine fire and spontaneous combustion.Since the development of gateroad has just completed, and retreat mining has not begun, the effectiveness of the proposed methods is unknown at this point. However, monitoring will continue until after mining. The results will be published in a separate paper.INTRODUCTIONGeneral descriptionTashan mine is located about 15 Km south of the city of Datong, Shanxi province, China (aure 1). The 2014 raw coal production was 25 million mt."

Jan 1, 2015

By Hutchinson IN

In order for the Australian underground coal mining industry to achieve the high productivity rates demanded of it without compromising its good safety record, improved geotechnical monitoring systems are required. One of the main problem areas affecting underground coal mining is the deterioration of previously stable roadways.CSIRO Division of Geomechanics has been working on a NERDDC funded project to develop an instrument to monitor coal mine roof deformation. The instrument which has been developed, Roof Alert, is a self contained, battery powered, intrinsically safe device. It can monitor the magnitude and rate of deformation and give multiple, variable visual alarm signals where preset levels ate exceeded.

Jan 1, 1991

By Norbert Patzelt, Hans Knecht, Wolfgang Baum

Future ore reserves of major gold producers in North America and elsewhere will consist of refractory gold ores. About 60 % of the current U.S. gold production is mined in Nevada. Several of the large gold operations in this state indicate that 40 to 90 % of their future ore feeds will contain refractory gold. To compound problems, average head grades for refractory U.S. gold ores have fallen from 8 - 9 g/mt to about 4 g/mt by 1991. In Latin America, 52 % of the gold operations will require milling for optimal recovery. Deeper reserves are, like in North America, refractory. Particularly in hard, (semi)refractory sulfide ores, many operations may experience substantial decreases of gold recov­ery. For example, Martins et al. (1991) report that at an Austra­lian mine, sulfide ore treatment led to 4 - 10 % lower gold reco­veries as the work index increased from 8 kWh/t to 14 kWh/t and the grind coarsened. Many refractory ores should actually be classified as semirefractory since a fine sulfide fracturing and liberation would make 80 - 90 % of the gold cyanide leachable. However, conventional comminution is often prohibitively expensive for these ore types.

Jan 1, 1995

By AIME AIME

HENRY YONGE WALKER is one of Canada&apos;s numerous gifts to the American mining and metallurgical industry, having been born it1 New Brunswick 59 years ago. At eighteen he came to the United States and has since been continuously identified with what are known as the Guggenheim interests. His first position was in the capacity of clerk at the Philadelphia Smelting & Refining Co. plant of M. Guggenheim Sons at Pueblo, Colo. In 1905 he was transferred to the plant of the American Smelting & Refining Co. at Everett, Wash., as assistant manager, two years later going to the Tacoma plant in the same capacity though with added

Jan 1, 1938

By Syd S. Peng

Introduction Modern longwall mining, introduced to the US coal industry in the mid-1960s, is the latest coal mining technique. Today, longwall mining produces more than 15% of all underground coal production. The growth of longwall mining in the US is slow. However, nearly two decades of longwall mining have demonstrated that its benefits outweigh the disadvantages. Longwall Advantages Production and Economics - A typical longwall production ranges mostly from 0.9 to 1.8 kt (1000 to 2000 st) of clean coal per shift. This is about three to six times the production of comparable continuous mining units. Furthermore, there are much less rejects in longwall mining, typically 15% to 25%. It must be noted, however, that a new longwall with an inexperienced crew may produce less, depending on mining conditions. But as experience gains, production increases rapidly. Due to its high production, the cost per ton of coal mined over the mine&apos;s lifetime is cheaper with longwall mining than with continuous mining. A study on a conceptual mine of 25 year life using actual operational data (Dangerfield, 1981) concluded that the cost per ton of coal mined is 32% cheaper by longwall than by continuous mining. Another benefit from its high production potential is that it enables the production to be concentrated in a few longwall faces versus many units of continuous mining required to achieve a similar production. As a result, mine organization and management simplify considerably. Safety and Savings - Most US longwalls use shield supports that cover the roof with solid canopy and isolate a gob completely from the face area. Thus, if the supports are properly selected and operated, the potential for roof fall accident is almost zero. Furthermore, it is not unusual that a well-run longwall face is much more orderly and cleaner that most well-run machine shops. The ventilation system for longwall mining is the most ideal type. It is simple and unique. No auxiliary fans are required. The fresh air sweeps through the whole face. Its volume and velocity can be adjusted as demanded. Recovery - In spite of multiple entries layout for US longwall panels, the longwall method recovers up to 40% more coal than the continuous mining method (Dangerfield, 1981). The typical recovery for longwall mining is 70% to 85% versus 45% to 80% for continuous mining. Versatility - Longwall mining has been employed successfully in seam heights ranging from 0.8 to 3.7 m (2.7 to 12 ft); in seam inclinations from horizontal to 35°; in overburden depths from 46 to 914 m (150 to 3000 ft); in single or multiple seam mining; and in uniform or irregular seam characteristics. For seam height less than 1.2 m (4 ft), the plow is used instead of a shearer. However, in spite of recent developments, the plow still is unsuitable for hard coal seams. Slow Acceptance How come longwall mining, with so many advantages, has not spread more rapidly? Undoubtedly, the major reason is the large capital investment required. Under normal market conditions, the capital required for a longwall panel ranges from $7 to $11 million. Another factor is the uncertainty of its applicability to any specific coal seam of interest where longwall has not been used or operated successfully. Although longwall mining elsewhere has proven its applicability to wide-ranging seam conditions, there have been several failures in the past two decades. Analyses of those cases show that most failures can be attributed to inexperience, both operational and technical, including equipment selection. Therefore, personnel training on the longwall mining technique is absolutely necessary before its initial operation. Some earlier statements about adverse surface subsidence associated with longwall mining were misleading. Research has proven that surface subsidence under longwall mining is controllable and predictable. And, if longwall mining was mandatory since the onset of coal mining in the US, there would be no surface subsidence problems over abandoned mined land today. It is a blessing that surface subsidence occurs

Jan 3, 1985