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By G. L. Boxer, R. A. Metz, C. B. Smith, W. J. Atkinson

Since the inception of the Ashton (formerly Kalumburu) Joint Venture in 1972 the parties have expended $24 million in exploration for diamonds in the Kimberley region. This has resulted in the discovery of the currently subeconomic Ellendale deposits in 1976/77 and the economic Argyle - Smoke Creek Deposits in 1979/80. In addition to Argyle and Ellendale, some 90 kimberlitic and lamproitic bodies have been discovered grouped broadly into three provinces, on the west, east and north margins of the Kimberley block. These bodies range in size from dyke-like features less than 1. m (3 ft) in width to pipes with a surface area of 128 ha (316 acres). The bodies with larger surface area are volcanic crater deposits, infilled with air-fall and water deposited tuffs and epiclastic sediments. Diamond content ranges from trace to economic concentrations. The exploration discoveries were facilitated by early recognition that the petrography, mineralogy and chemistry of the diatremes varied from the classical kimberlites of Kimberley, South Africa, to unusual more highly fractionated, alkaline, silicic varieties having affinities with leucitelamproite. Indirect methods of exploration were used, primary reconnaissance work being stream gravel sampling for kimberlitic indicator minerals. Evaluation showed Ellendale diamonds are 90% gem and near gem, but the low grades and tonnage render the deposits unpayable at today's diamond prices. After discovery in late 1979 and completion of a feasibility study by the end of 1982, evaluation studies proved Argyle to be economic. Alluvial diamond mining commenced in 1983 producing 6.2 million carats (Mct) that year, 60% being of industrial quality. Open-pit mining of the AK1 lamproite should start in 1985/1986 at a rate of 3 million t/a (33 million stpy) and produce 25 Mct/a, making this mine, and Australia, the world's largest producer of diamonds by weight. The revenue from diamond deposits is a function of the grade and the value of the diamonds. The grade is the product of the frequency of occurrence of the diamonds (stone density) and the mean stone size; the value of the stones is strongly related to their size distribution. During evaluation of the West Australian deposits, the diamond distribution was found to vary markedly in stone density and stone size. These variations corresponded with changes in the rock type hosting the mineralization. As a consequence, geological mapping of individual lamproites within the multiple vents and careful monitoring of the diamond distribution became an important part of the evaluation work. Similar variations in stone size and frequency were found to accord with separate alluvial terraces and fan gravels derived by erosion from the Argyle AM lamproite.

Jan 1, 1992

By James F. Howe

EveRy engineer and user of wire rope is desirous of information that will enable him to determine whether the performance of any particular piece of rope is satisfactory, and what conditions can be changed so as to increase the life of a rope. In most cases, it is far from easy to analyze the conditions of service and tell which is the greatest factor in determining the life of wire rope. If we start on a layout for an entirely new rope system, the task of making conditions satisfactory is simplified so far as the engineering details are concerned; but if the plant requiring wire rope is in operation, any change is likely to involve the expenditure of considerable money to replace and change over machinery that must be kept in constant operation, because a large monetary loss will result from a shutdown for even a few days. Often wire rope is kept operating under bad conditions, as the user thinks that it is cheaper and quicker to replace the rope than to make alterations in the hoisting machinery. A typical case is a metal mine operating in what is now almost the center of a large mining city, where it owns a few acres of mineral rights. It was expected that the shaft would go to a depth of 1000 ft. (305 m.), so a moderate-sized rope, about 1 1/4in. (31.7 mm.), was used on a 6-ft. drum with a 7-ft. head-frame sheave and rather short lead from drum to head sheave. This mine has gone deeper each year until it is now 2500 ft. deep and the rope speed has been increased from 1500 ft. per min. to over 3000 ft. The original rope just filled the drum in one layer, now it overwinds three layers. The original ropes were probably 6 X 19 crucible steel and lasted from 12 to 18 mo.; now, the highest grade plow steel is used, yet the average service is only about 6 mo. The cause is not hard to find. If the engineers had known at the outset what would be done with this hoist, the design would have been very different. Conditions at this mine, from a wire-rope standpoint, have gradually become worse and it is difficult to see how rope service on this hoist can be compared with the service of previous years with any fairness to the wire-

Jan 1, 1923

By James F. Howe

EveRy engineer and user of wire rope is desirous of information that will enable him to determine whether the performance of any particular piece of rope is satisfactory, and what conditions can be changed so as to increase the life of a rope. In most cases, it is far from easy to analyze the conditions of service and tell which is the greatest factor in determining the life of wire rope. If we start on a layout for an entirely new rope system, the task of making conditions satisfactory is simplified so far as the engineering details are concerned; but if the plant requiring wire rope is in operation, any change is likely to involve the expenditure of considerable money to replace and change over machinery that must be kept in constant operation, because a large monetary loss will result from a shutdown for even a few days. Often wire rope is kept operating under bad conditions, as the user thinks that it is cheaper and quicker to replace the rope than to make alterations in the hoisting machinery. A typical case is a metal mine operating in what is now almost the center of a large mining city, where it owns a few acres of mineral rights. It was expected that the shaft would go to a depth of 1000 ft. (305 m.), so a moderate-sized rope, about 1 1/4in. (31.7 mm.), was used on a 6-ft. drum with a 7-ft. head-frame sheave and rather short lead from drum to head sheave. This mine has gone deeper each year until it is now 2500 ft. deep and the rope speed has been increased from 1500 ft. per min. to over 3000 ft. The original rope just filled the drum in one layer, now it overwinds three layers. The original ropes were probably 6 X 19 crucible steel and lasted from 12 to 18 mo.; now, the highest grade plow steel is used, yet the average service is only about 6 mo. The cause is not hard to find. If the engineers had known at the outset what would be done with this hoist, the design would have been very different. Conditions at this mine, from a wire-rope standpoint, have gradually become worse and it is difficult to see how rope service on this hoist can be compared with the service of previous years with any fairness to the wire-

Jan 1, 1923

[Holds monthly meetings, except June, July and August, J JOHN A. CHURCH, JR., Chairman JOSEPH E. POGUE, Vice-chairman SIDNEY ROLLE, Vice-chairman EDWARD C, MEAGHER, Secretary, Texas Gulf Sulphur Co,, 41 E, 42nd St., New York M, H, MERRISS, Treasurer, 42 Broadway, New York ARTHUR NOTMAN T. T, READ Boston Meets first Monday of each winter month. EDWARD C. GEITHER, Chairman WILLIAM S. BLACK, Vice- chairman DONALD H, MCLAUGHLIN, Secretary, treasurer, Rotch Bldg,, Harvard Univ,, Cambridge, Mass. J, L. GILLISON A. G, WAITE]

Jan 1, 1928

By Walter L. Maxson, Fred C. Bond

SINCE the last publication of tabulated results of grindability tests by the authors1 the total number of ball-mill tests made has more than doubled, and rod-mill tests have become increasingly important. Nearly all of the standard closed-circuit ball-mill and rod-mill tests made to date are included in the present tabulation, together with additional information not previously published. STANDARD BALL-MILL GRINDABILITY TESTS The method of conducting these closed-circuit tests has not been altered. A sample of the ore or other material is stage-crushed in rolls set at 1/16-in. opening with a 6-mesh screen until all of it has passed the screen. The combined minus 6-mesh screen undersize is mixed, sampled and screen-analyzed, and its apparent specific gravity is determined by packing and shaking in a standard container, and weighing. The apparent specific gravity ordinarily is about 60 per cent of the true specific gravity. The unit volume present in the mill in all tests is 700 c.c. of the packed minus 6-mesh roll product, and the number of grams occupying 700 c.c. is the unit test weight. This weight is placed in the mill dry, ground for the number of revolutions estimated to be necessary, discharged, and screened mechanically in three testing sieves of the mesh size at which the test is to be conducted. The oversize, or circulating load, is weighed, sufficient fresh minus 6-mesh feed is added to bring the total weight up to that of 700 c.c., and the charge is returned to the mill for a second grinding period. All standard ball-mill tests are conducted at 250 per cent circulating load, and the number of revolutions in the mill necessary to obtain this circulating load at any grinding period is estimated from the results of the previous period. The number of net grams of screen undersize produced per revolution of the mill approaches its final equilibrium value after several grinding periods, and this is recorded as the grindability, or relative ease of grinding, of the sample. A cylindrical ball mill, 12 by 12 in. inside, with a smooth lining and rounded corners, is used with a revolution counter. Under standard grinding conditions it runs at 70 r.p.m., and contains a charge of 285 iron balls weighing 20,125 grams, ranging from I ½ to ¾ in. in diameter. Tests are conducted at the mesh size to which the ore is to be ground in practice, and the capacity to be expected from a given mill is calculated by comparing the

Jan 1, 1943

By C. E. Nighman, R. S. Foster

The following is a description of the methods used in rescuing men and extinguishing the underground fire at the Pennsylvania mine, Butte, Mont. , This fire, which cost the lives of 21 men, began about 9 p.m., Feb. 14, 1916, at or near a ventilating fan on the 1,200 air-shaft station. This shaft at that time was downcast, while the main hoisting shaft was upcast. Owing to the proximity of the fire to the air shaft, the air in the latter soon became heated, and as the fan on the level was not running— the power having been cut off shortly after the outbreak of the fire—the air shaft became upcast to the 300 level. At this point a blanket of cold air from the surface caused the smoke and gases to be drawn in on that level and carried down again through the workings to the 1,200. A short time after the discovery of the fire, water was turned down the air shaft through a series of pipes around the collar of the shaft, which caused the smoke to recede below the 300, so that within 2 hr. it was possible to search that level for the missing men. From the above it will be apparent why the air shaft was upcast instead of downcast, as under normal conditions, and why men were suffocated at considerable distances from the seat of the fire. The smoke and gases that came up the main hoisting shaft did not diffuse through the workings to any extent. There are doors on all levels in the connections to the main hoisting and air shafts, but in the confusion and running around many of these doors were left open, and this undoubtedly caused the smoke and gases to spread more rapidly than they would otherwise have done. It was necessary for most of the men on the various levels to pass close to the air shaft in going to the hoisting shaft, and when they encountered the smoke and gases in that vicinity a number turned back and climbed down with the direction of the air currents and were then suffocated, while all those who passed through the pall of smoke to the main shaft were saved. There were 220 men in the mine that night, and all but 25 were hoisted to the surface within 30 min. after the discovery of the fire. Of the 25, five went out through the Tramway mine, and one through the Moun-

Jan 1, 1918

By Carla Lupi, Mauro Pasquali

"Photographic technology is based on light sensitivity of silver halides that, because of silver properties, are capable of forming photographic images. Photographic laboratories and hospital departments because of digital photographic systems less use this technology today. Nevertheless recovery of silver is still advantageous because environmental problems related to Ag ions danger to biological system. Many processes have been studied to treat silver containing photographic bath: cementation, chemical precipitation and electrowinning. In this work, after preliminary voltammetric tests, a batch electrometallurgical process has been studied to completely deplete the silver containing photographic fixing baths. The effect, of silver concentration in the electrolyte, varying over the range 0,2-1 g/L, bath stirring and current density ranging from 10 to 500 A/m2 on process yield and deposit quality, has been investigated. Introduction Photographic wastes, electronic scrap and spent catalysts constitute the main sources of materials for silver recycling. Among those silver secondary sources, the photographic discharged material could be still important and useful, even if photographic laboratories and hospital departments, because of digital photographic systems, less use photographic technology. Several processes have been developped in order to recover silver used in the field of photography from both photographic solutions and scrap film paper[1-5]. Fixing solutions are very complex baths but, because of silver position in the electromotive series, electrolysis could represent an advantageous method for silver recovery. Electrolysis produces high purity silver without generating new pollutants, while the fixer can be reused: obviously it is necessary to purge off a portion of this solution in order to avoid the excessive build up of other deleterious compounds into solution. A tested recovery method was constituted by the dissolution of silver from X-ray films with concentrated nitric acid. Internal electrolysis and external electrolysis methods were carried out using a steel electrode as cathode and a copper electrode as anode [6]. Silver recovery from photographic chemical solutions has been also performed in a special cell [7]: the solution is supplied to the bottom of a non-agitated cell having a vertically elongated form and including a central anode enclosed by a cathode. An other study discloses the use of an electrolytic cell comprising at least one anodic compartment and one cathodic compartment: the solution is treated in the cathodic compartment and then recycled in the anodic one [8]. A more recent work proposes the use of zirconium cathode for the recovery of silver from photographic fixing solution [9]: studies carried out by cyclic voltammetry showed that about 98% of total Ag can be recovered with an electrical efficiency close to 92%."

Jan 1, 2008

Potential hlghwaJl mining reserves in Australia are large, but geotechnical and operating conditions are quite different from those where highwall mining has evolved in the USA. In order to achieve safe and effective resource recovery, equipment and procedures should be developed to penetrate 250 m in the Australian mining environment. This paper identifies key auributes to be incorporated in such machinery and methodology. These suggestions are intended to promote discussion on innovations to satisfy Australian requirements for highly productive, safe and environmentally acceptable techniques.

Jan 1, 1993

By Carlos Rebolledo, Nelson Parra

Arsenic is an unresolved problem in the mining industry considering the fact that most of the new copper deposits have high arsenic contents; stricter environmental restrictions for handling, transporting and processing complex materials are expected and stable abatement of arsenic is not extensively applied. In 2018 Chile produced 24.5% of the total copper from concentrates and in the future copper production in Chile is expected to increase, mainly associated with production of concentrates that will also contain more arsenic. China has set an import benchmark of maximum 0.5% for arsenic content in copper concentrates and this value is expected to be more restrictive in the future. On the first day of 2018 China banned the importation of scrap. Blending practice has been a business option, but it is not sustainable in the future, as “normal” copper concentrates become less and less clean. By the end of 2018, Chile formulated new regulations, increasing the capture of SO2 and arsenic to 95%. In order to comply with these regulations Chilean smelters had to make large invest large amounts of money. Disposing of arsenic as a stable residue is a pending issue. EcoMetales is a 100% subsidiary of Codelco and provides the mining industry with environmental solutions. It has a plant near Calama in Chile that treats flue dust from smelters, refinery effluents and acid solutions from acid plants. EcoMetales is fully committed to contributing to the solution to the arsenic problem. Currently we are developing four main projects: 1) Concentrates treatment. A PLCC project considering the treatment of 200 ktpy of copper concentrates with arsenic from the northern district of Codelco; 2) EPAS DET project considering the treatment of 700 m3/d of weak acid solution with arsenic from the of El Teniente devision of the Codelco acid plant; 3) Borras project, considering the recovery of valuable metals (Cu, Ag, Ge, Pb) from the flue dust leaching plant; 4) Tailings project, considering the recovery of metal values from tailings. The purpose of this work is to present the trends and treatment processes for arsenic handling in the copper mining industry with a focus on the scorodite process. INTRODUCTION Between 2000 and 2015 the global copper concentrates production has grown from approximately eleven to fifteen million metric tons per year. Chile produced 26.5% of the total copper from concentrates and in the future copper production in Chile is expected to increase, mainly as a result of the production of concentrates (Figure 1)

Jan 1, 2019

When most people think of coal mining, they envision men in soot-covered clothes and hard hats digging at deposits in deep underground caverns. These mental images commonly include coal cars rumbling on rails as they emerge from the earth, filled to the brim with rocky ore. The world of modern, longwall coal mining is a far different place. Yes, it?s underground, the miners still wear hard hats and the environment is filled with soot. However, today?s mining operations are carefully monitored and the entire process is highly automated. Five kilometers (3 miles) underground, a sophisticated, high-powered mining machine moves steadily through a 2-m (6-ft) high tunnel at an average speed of 3.2 km/h (2 mph), carving out coal, then conveying it to a crusher and carrying it to the surface. Instead of digging out the ore, a longwall mining system slices it from the exposed tunnel face, using giant rotating shears. The shorn coal tumbles onto a conveyer below the shears and is conveyed to the stageloader, where it is crushed and dumped onto the main conveyer belt that eventually hauls it out of the mine. Integral to the longwall shear and conveyor is a moving metal roof that protects it from the thousand plus feet of ground above. The ceiling panels automatically advance above the shears, powered by hydraulic roof supports that stand on the other side of the tunnel, moving in a slow, sideways lock-step with the shears. This complex and massive machine makes longwall mining the most ef?cient and productive method for excavating coal. In fact, it produces 2.7 kt to 22.6 kt (3,000 to 25,000 st) of ore per shift, versus 544 to 907 st (600 to 1,000 st) for the conventional room-and-pillar mining. Using that method, miners develop one section of a mine at a time after carving out load-bearing columns of coal to support the roof above their heads. Longwall mining also is considered safer than the room-and-pillar method, largely because the process is mostly automated, requiring fewer workers. Still, longwall mining has its drawbacks: The equipment is expensive and the mining conditions are harsh. An entire longwall system can cost between $50 million and $100 million, while the electrical controls used to operate it can cost an additional $1 million to $2 million. These investments grind away all day, every day, in hot tunnels filled with coal dust.

By Alwyn Protheroe, Steve Smith, Herbert Jr. Gibson

In September of 2000, Cowin and Company, Inc., contracted with Paramont Coal Corporation to construct a large decline approximately 820 m (2,700 feet) in overall length at a grade of 8 degrees to the Lower Banner Coal Seam. This paper will discuss the methods used by the contractor and owner to overcome the unique problems required to construct this tunnel in a coal mine environment. The problems include, initial design considerations, equipment selection, ventilation, ground control, control of water in?ows, and placement of concrete.

Jan 1, 2003

By Steve Kral

"Gold rush" is usually associated the western United States - California, Nevada, and Colorado. The southeast section of the nation is known more for its rice, textiles, tobacco, indigo, and cotton. This area, though, is where it all began with gold in this country. The United States' first gold rush, in fact, is said to have started with the discovery of a 7.7-kg (17-1b) nugget in the Piedmont section of the Carolinas in 1799, some 50 years before the rush to California. The Carolina gold rush continued intermittently through 1942 when the Federal Government shut down all gold operations. Strong gold prices throughout the early part of this decade and new extractive technologies spurred a small resurgence of gold operations in South Carolina. Mind you, this is not Nevada with its flood of geologists and mining companies. Nor are there any 7.7 kg (17 lb) nuggets being discovered. But Ridgeway Mining Co., a joint venture between Kennecott Corp. and Galactic Resources, has set about mining 51 Mt (56 million st) of low-grade ore containing an estimated 43.5 t (1.4 million oz) of recoverable gold. The open pit, carbon-in-leach operation is located near the town of Ridgeway in South Carolina, about 32 km (20 miles) north of Columbia. Carolina gold North and east of Ridgeway, Piedmont Mining Co. operates the historic Haile gold mine near Kershaw, SC. And Brewer Mining Co. operates the nearby Brewer deposit. Between 1829 and 1942, South Carolina gold mines produced 10 t (318,800 oz) of gold. It all began in 1799 when Conrad Reed, son of a German immigrant farmer, found the 7.7 kg (17 lb) nugget in Little Meadow Creek, NC. His father thought the rock was worthless and used it as a doorstop, so the story goes. It took nearly three years before Conrad's father decided his doorstop was indeed gold. Gold was later discovered in the Greenville district of South Carolina. And, in 1828, Benjamin Haile discovered gold in a stream that ran through his property in Lancaster County, SC. Haile's mine became South Carolina's most famous and most productive gold mine, producing more than 7.8 t (250,000 oz). At first, miners worked the stream beds of Haile's property. Later, lode deposits were discovered on slopes alongside the creeks. They would lease 4.6 ml (50 sq ft) lots and mine as deep as 7.6 m (25 ft). During the Civil War, the Confeder-

Jan 1, 1989

Members of Fifty Years or More Lawrence Addicks '02 T H Aldrich '98 Walter H Aldridge '89 Eugene Ant. '99 George H Ashley '99 R M Atwater, Jr '97 James B Bailey '00 Samuel Barker, Jr '00 John F Berry '93 Charles A Bohn '93 W S Boyd '02 H C Boynton '01 H A Brassert '99 Arthur C Brinker '01 James F Brophy '01 Davenport Brown '02 Donald C Brown '01 Victor B Buck '98 Frederick Burbidge '87 Frederick P Burrall '97 John Cabot, Jr '99

Jan 1, 1952

By John M. Hird

A section of unstable overburden that has hampered an all-dragline mining operations at Texasgulf's phosphate mine in eastern North Carolina, is being successfully removed by hydraulic dredging. Blocks are dredged, drained, and developed ahead of two large draglines that strip the remaining overburden and mine the 12-m section of ore. Dredging assistance has increased ore production, increased ore recoveries to a maximum, reduced unit costs, and has made it possible to reclaim mined-over land that would be left underwater mined by all dragline methods.

Jan 1, 1981

By M. M. Champion

INTRODUCTION Since the first introduction of explosives into mining operations, safety considerations have prohibited the use of machinery in association with the explosives. Only within the last 25 years has this situation changed appre¬ciably; the closest thing to "machinery" used by a powderman in the "old days" was his wooden tamping stick. With the blackpowder and dynamite used during the first 150 or so years of mining blasts, substituting a drill steel or a metal "spoon" invited serious disaster. With the advent of ammonium nitrate-fuel oil (ANFO) explosives (1950s) and water-gel explosives (1970s) for underground blasting, machinery started to play a large part in blasting procedures. The lower sensitivity of ANFO and water gels allows these prod¬ucts to be loaded in a variety of safe ways. Not only has this meant cost savings in labor and blasting materials, but it has also resulted in whole new mining concepts where the limitations of dynamite previously made those concepts uneconomical, impossible, or excessively risky. Increasingly (totally for some mines), the traditional tamping stick has given way to a variety of pneumatic machines for loading explosives and blasting agents. The resulting blasting operations have become safer, more efficient, less uncomfortable for personnel, and a major step forward in mining technology. ANFO LOADING EQUIPMENT Three basic types of loaders are used for ANFO explosives, these being the pressure-vessel loaders, Venturi loaders, and combination loaders. Pressure-Vessel ANFO Loaders ANFO consists of prilled ammonium nitrate (porous pellets I to 2 mm in diam) and fuel oil in the appro¬priate 94 to 6% blend. The material is fairly dry, pourable, and easily conveyed by an air stream. One device or system for handling (hole loading) ANFO is the pressure-vessel loader. ANFO is dumped into the vessel or "pot," and the vessel then is pres¬surized. The ANFO is conveyed out of the bottom of the vessel into a semiconductive hose that has been inserted into the blasthole. Fig. 1 illustrates a typical pressure-vessel loader in use. Applications: The use of pressure-vessel loaders is widespread in room-and-pillar mines, including salt, potash, trona, lead, and other rubber-tired mines where the heavy pressure vessels can be mounted on a vehicle for transport from face to face. The vessel often is large enough to hold eight or more 23-kg (50-1b) bags of ANFO, so several faces can be loaded before refilling. Pressure-vessel loaders also are suitable for effec¬tively loading long holes such as those that are 6 m (20

Jan 1, 1982

By Shao Li, Chen Zhiming, Wu Jihong

In this paper,powder emulsion explosives,which have high sensitivity and strong detonation force,have been studied by applying emulsion technology to the production of the present powder ones.Under the condition of #8 industrial detonator initiation,the distance of air gap propagation is 110~180mm,brisance !3.5mm, lead block test value 325ml and detonation velocity 3000~4000m/s for small diameter (35mm) explosive cartridges.In addition to these,the factors effect on are investigated by experimental study and theoretical analysis.

Jan 1, 1991

By Clive Wilson, Bruce Hutchison, Daniel N. Adams

Two three-lane highway tunnels, 700 and 2900 m long, make up the driven tunnels portion of Melbourne City Link. Tunnel construction, using design-construct delivery, commenced in May 1996. Construction required excavation through varying rock conditions; temporary support included ground freezing, pipe-spiling, steel sets, and rock bolts. Permanent support for the tunnels required both drained and undrained tunnel cross sections using unreinforced concrete. This paper reviews the construction methods, and highlights some of the not-so-standard designs developed while construction was underway.

Jan 1, 1999

In April, 1990 the Sydney Water Board invited bids for the construction of the Hazelbrook Carrier Section 3 being part of the Blue Mountains Sewerage Transfer Scheme. In December 1990 the contract was awarded to Costain Resources Limited, Mining Services Division for the sum of $A15.3 M. The works consisted of excavation of 9 520 linear metres of tunnel, 3.35 metres diameter with installation of permanent ground support in the form of rockbolts, mesh, steel sets, shotcrete and in situ concrete

Jan 1, 1993

By H. P. Carlisle

The significance of the mineralogical and surface characteristics of gold grains in the recovery process by C. E. FEATHER and G. M. KOEN Published in the Journal. February 1973. DISCUSSION H. P. Carlisle. B.Sc. (Eng.) Rand (Fellow) The authors are to be congratulated on drawing attention to the importance of surface characteristics in the treatment of gold grains. I must, however, confess to considerable doubts about the validity of the distribution of gold in "free" and "locked" form set out in Tables I to IV.

By Michael E. Lively

DOUG HALBE: We've talked about copper operations and gold operations. For our third section on the topic "Operating Ideas That Have Saved Me Some Bucks," we'll turn to the coal fields. Our featured speaker is, I'm almost embarrassed to say, a mining engineer, Mike Lively. Mike has 25 years' experience in the coal industry. He is currently superintendent of surface operations for Southern Ohio Coal Company, Meigs Division (AEP Company). I've read his paper, and there's some really great information here. Mike.

Jan 1, 1998