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By Lauren M. Roberts, Ralph R. Sacrison

The Meikle Mine uses long-hole open stoping and cut-and-fill methods with backfill as the primary ground support technique. The mine operates at 2,858 tonnes per day (3,150 tpd) and has sustained operation in excess of 3,175 tonnes per day (3,500 tpd). Consistent backfill production is critical in achieving these mining rates. The heart of the system is a 7.65 cubic meter (10-cyd) horizontal shaft ribbon mixer located 328 m (1,075 ft) underground. System control is through several PLC's, allowing for a high degree of automation and process flexibility. Aggregate feed is routed 257 m (842 A) through vertical drop pipes in the ventilation shaft, a rockbox, transfer raises, and then into either of two raisebored storage silos. The aggregate is crushed mine waste, nominally 70% 3.8 cm x 1.0 cm (1-1/2"x 3/8") and 30% -1.0 cm (- 3/8"). The binder is routed through a vertical borehole, surge bins, and transfer raises. The binder is a mixture of Type II cement and fly ash. Two primary manufactured fill designs are produced: a high strength mix for stopes which will be undercut; and a low strength mix for stopes with only rib exposure. The high strength target is 5.5 Mpa (800psi) unconfined compressive strength in test cylinders. The low strength target is 2.8 Mpa (400 psi). The systim has allowed consistent achievement of these targets, while experience has allowed reduced costs and increased efficiencies. Fill placement is through 6.1 cubic meter (8 cyd) LHD's dumping into stopes. Specific sills and cut-and-fill drifts are jammed to the back. System enhancements have included colloidal mixers for cemented waste slurry placement and transfer raises for sending manufactured fill to the lower levels of the mine.

Jan 1, 2001

By Anna Kate Miller, James J. Brady, Hamza Jaffal, Ritika Sangroya, Chadi El Mohtar, Wing Shun Kwan

"The design of an appropriate pre-grouting program can be the difference between a successful tunnel construction and an expensive on-the-go repair. However, current grout design relies heavily on rulesof- thumb and outdated charts. The geotechnical investigation is rarely targeted towards grouting design and the final grouting decisions are made based on fines content or soil classification. The current paper presents an experimental study performed to determine the optimal grout for a tunnel in soft ground below the ground water table (GWT). The testing program consisted of sieve analysis and undrained static triaxial tests on reconstituted sonic samples. Rheological tests were used to compare the different grout mixes and identify best candidates before performing one-dimensional permeation tests. Pressure and volume measurements during the permeation tests, along with the unconfined compression strength results on the permeated soils, helped evaluate the filtration potential for the different grouts and select the optimal grout.INTRODUCTIONConstruction of the Millwoods Double Barrel Replacement Project commenced in 2009 and was expected to be completed by 2015. To facilitate ongoing city development in south Edmonton, the City implemented construction of a major sanitary trunk network along 91st Street and 30th Avenue. The project involved the design and construction of approximately 1.2 miles (2 km) of 7.8-ft (2,400 mm) sanitary tunnel and 2.1 miles (3.4 km) of 11.5-ft. (3,500 mm) and 9.5-ft. (2,900 mm) storm tunnel. Part of the project also involved the conversion of the existing double barrel storm and sanitary trunk to full sanitary service, and reconnecting the local storm sewers from the double barrel pipe to the new storm tunnel. The tunnels range in depth from 26-ft. (8 m) to 115-ft. (35 m) and are being installed using a one pass system with precast concrete segmental liners."

Jan 1, 2016

By William T. Barrett

It is said that Thales of Miletus, 600 B.C.,1 was aware of the attraction of lodestone toward iron and highly magnetic minerals. However, the birth of magnetic concentration followed shortly after the invention of the first electromagnet by Sturgeon in 1825. Sturgeon's magnet was described in the annals of electricity.2 The first attempts to concentrate minerals by the use of magnets date back to about 18473 but the first commercial drum-concentrator was apparently invented by R. Cook in 1849. Credit must also be given to the invention of P. Righter's machine in 1867,5 and to the highly successful Wenström separator devised in Sweden in 1883.8 Since the advent of these early separators there have been virtually hundreds of patents granted covering minor modifications of drum-type magnetic separators. The more important advances from a technological viewpoint have been the design of stable power supplies for d.c. magnets, the advent of ceramic permanent magnets which are now replacing electromagnets for the concentration of ferromagnetic minerals, and advances in general magnet and machine configurations.7 One of the problems facing both the designer and the user of magnetic separators is the evaluation of the body forces acting on particles in the magnetic field of the separator. Perhaps the most elegant method of evaluating the body force associated with a drum-type magnetic separator is that recently suggested by W. F. Brown.8 Brown measures the radial component of the field intensity one-half inch from the drum surface and fits a Fourier series in the angle about the drum axis (Figure 1). Once the coefficients of the series are known, it is then possible to calculate [ ]

Jan 1, 1970

By Anibal Gajardo

Because of the Chilean Ministry of Mining's interest in the development of industrial minerals, the Chilean Geological and Mining Service (Servicio Nacional de Geología y Minería, SERNAGEOMIN) has carried out a research program during the last two years to identify projects that represent interesting investment opportunities for small- to medium-scale operations in Chile. Most important of these projects are: (1) Quebrada Lluta aluminum sulphate, for drinking- and industrial-water treatment, with a potential domestic and foreign mar- ket of almost 43,000 tonnes per year and investments of about US$1.8 million; (2) Quebrada Lluta-Quebrada Cardones pumice, for abrasives, filters, fillers, and light- weight aggregates, with a potential market of about 30,000 tonnes per year, and in- vestments of US$2.0 million; (3) El Maule dolomite, for soil conditioner and iron smelting, with a total potential domestic market of some 60,000 tonnes per year, and estimated investments of US$2.0 million; (4) Las Escaleras bentonite, for animal feed, pelletising, and oil absorbents, with a potential market of almost 6,000 tonnes per year, and investments of about US$2.4 million; (5) Laraquete-Carampangue andalucite, for alumina refractories, with a future total domestic and foreign demand of about 15,000 tonnes per year, and estimated investments of US$2-2.5 million; (6) Rio Lía garnet, for fillers and sand-blasting abrasives, whose domestic market represents 10,000 tonnes per year, with investments of US$2.3 million; and (7) Munilque pumice, for building bricks and panels, with a potential domestic market of 250,000 tonnes per year, and investments of US$1.4 million. The development of these projects will represent an important increase and a significant diversification in Chilean industrial-minerals production.

Jan 1, 1999

By Richard R. Radel, Jack Barry

Contract required excavating 2,000 If of 14' horseshoe tunnel in rock, a 95' deep work shaft on shore, and a 50' deep intake shaft located in the middle of the Niagara River just above Niagara Falls. Large quantities of H2S were encountered in the range of 20 PPM with ventilation. Peak levels of 40 PPM were recorded. The original designed ventilation system required only a volume of 10,000 CFM to meet OSHA requirements. The volume of air was increased to 50,000 CFM to dilute the H2S to al- low workmen to work and comply with OSHA H2S action levels. The second phase of the tunnel consisted of placing 72" ID X 20 foot long concrete reinforced pipe in the tunnel. When the tunnel was holed through and both shafts were open, a substantial amount of work was required to take place in the shafts which eliminated the possibility of placing a bulkhead to secure proper ventilation. The ventilation team designed a jet system airlock. The design of this system was to bring fresh air down the work shaft in two duct systems; one 30" diameter and one 24" diameter. Following the excellent work of Jan Wolski (7th Mine Ventilation Symposium chapter 58) our design incorporated nozzles located approximately 100' in by the work shaft. The velocity of the 24" diameter duct was 6,365 FPMD2.2 MPH. The velocity of the 30" diameter duct was 6,117 FPMl69.5 MPH. Both systems were on blow and being that air expands at 3112 degrees per side this high velocity of air expanded and created an air bulk head through which men and the large sections of concrete pipe could be passed without any delay or additional cost to the project.

Jan 1, 1997

By L. Alan Weakly

This paper was presented at the ‘17th U. S. Symposium on Rock Mechanics’ on August 29, 1976, at Snowbird, Utah, at the SME-AIME Fall Meeting on September 1, 1976, at Denver, Colorado, and at the Polish-American conference on Ground Control in Room-and-Pillar Mining on August 7, 1980 at Southern Illinois University, Carbondale, Illinois. Introduction The purpose of this paper is to present a viable method of reinforcing rock pillars by utilizing fully grouted reinforcing steel and a cement- fly ash mortar. Although the fully bonded bolt concept is not new as exemplified by the present widespread use of resin bolts within United States mines and by the use of cement grouted reinforcing used as back pins in the Kiruna Iron Ore Mine in northern Sweden, (1)(2)the method of stabilizing mine pillars with fully bonded reinforcing is new. This method, called the "grout-rebar system", was first incorporated at St. Joe Minerals Corporation's Fletcher Mine in December of 1972. The system was used on an emergency basis before extensive research on the mechanics of the system could be completed. Since Decmeber of 1972, however, approximately 206 pillars have been reinforced at Fletcher and the system is currently being used on a regular basis at two other St. Joe mines and at two competitor mines in the area. Basically, this paper is a case study of the efforts, research and results of utilizing the grout-rebar system, on a selective basis, at the Fletcher Mine. In addition to information about the system, this paper will explain in detail the three rock mechanics instruments developed by St. Joe and used to measure and monitor rock movement to validate the group-rebar system.

Jan 1, 1982

By T. C. Page, P. C. Gibson, M. A. Miller, J. D. Sell

[The Yarnell gold deposit in the Weaver mining district, Yavapai County, Arizona, occurs in a structurally controlled, hydrothermally altered zone in a 1700 Ma granodioritic intrusive. Both potassic and sericitically altered rocks around and above the low-angle, northeast striking Yarnell fault are known to host economic gold mineralization. A wider envelope of weakly-propylitic altered rock also occurs in this area. During mineralization, strong sericitization accompanied several stages of quartz (±adularia) veining, stockwork formation, localized silicification and potassic replacement, in association with deposition of specularite, pyrite (now oxidized) and gold. The footwall of the fault is also sericitically altered but poorly mineralized. Mineralization along the Yarnell fault continues both northeast and southwest from the main deposit, although the thickness of the zone and associated alter avian envelope diminish away from the orebody A sample of undo formed illite taken from the Yarnell fault zone was K/Ar dated at 69 + 1.6 Ma. This date reflects a minimum age for both mineralization and latest movement on the fault. Gold mineralization is accompanied by modest increases in Ag, As. Cu and Mo content. Both CO2- and H2O- rich fluid inclusions were studied and suggest mesothermal pressures and temperatures of formation for the deposit. Ninety-six reverse circulation holes and four diamond drill holes outline a bulk-minable mineral reserve of'4.1 million tons at 0.051 oz of Au per ton. An additional 2.7 million tons at 0.017 oz of Au per ton occurs in a low-grade zone above and subparallel to the main mineralized zone. Total calculated mineral inventory stands at 6.8 million tons at a grade of 0.038 oz of Au per ton, with a waste-to-ore ratio of 1.45:1. Column leach tests indicate that cyanide heap leach gold recoveries should exceed 70%.]

Jan 1, 1994

By Wei Bin, Liu Zhe, Liping Xiao, Jichi Bai

The Acid Mine Drainage has characteristics of low pH, high concentrations of heavy metal ions, such as Fe2+, Mn2+, Cu2+ and Zn2+. In this paper, the composite particles that consisted of bentonite and steel slags were used to dispose the Acid Mine Drainage. Bentonite is the mineral materials and can adsorb heavy metal ions excellently. The steel slags are alkaline and industrial solid waste for the treatment of the Acid Mine Drainage. The main influencing factors, such as adsorbent dosages, shaking rates, concentrations of heavy metal ions, temperatures, adsorption time and pH value were studied by the static experiment. Increasing the adsorbent dosages and the concentrations of the heavy metal ions, speeding up the shaking rates, raising the temperatures, extending the adsorption time and improving pH would improve the removal efficiencies of heavy metal ions. Through considering the removal efficiencies and the treatment costs, the optimum reaction conditions for the four kinds of heavy metal ions were obtained. The adsorbent dosage 21 mg/L, the rotational speed 120 r / min, the temperature 25?, adsorption time 100 min and the initial wastewater pH 7. The four kinds of heavy metal ions existed independently in the simulated mine wastewater. The highest initial concentrations of Fe2+, Mn2+, Cu2+ and Zn2+ were 150 mg/L, 100 mg/L, 170 mg/L and 140 mg/L respectively. The removal rates were 93.42%, 92.64%, 93.86% and 95.17% respectively. The determination of the particles’ SEM-EDS Microscopic characterization showed the composite particles could play a part in neutralizing, absorbing and the chemical precipitation. The acidity decreased notably and the heavy metal ions of Fe2+, Mn2+, Cu2+ and Zn2+were adsorbed and precipitated effectively. The research results can extend the practical engineering application of the composite particles.

Jan 1, 2017

By J. Juilland

Over 10 percent of the United States of America or 41 percent of the Federal lands is managed by the Bureau of Land Management (BLM) as Public Lands. The mission is to manage the use of these lands and their resources. The way in which BLM carries out its mission is being refocused under the concepts of biodiversity and ecosystem-based management for the benefit of various human needs including environmental, social and economic values. The Bureau is developing methods to integrate the renewable and non-renewable resources programs into the concepts of biodiversity and ecosystem-based management. Exploration, development and extraction of energy or mineral resources on Public Lands are affected. The statements that following the next few pages are my personal views as a scientist (economic geologist -mining technologist) who is and has been involved for the last twenty years in the multiple-use management of Public Lands. They are not necessarily the position of the Bureau of Land Management, although much of the information on which I base my views comes from my experience with the BLM. The Bureau of Land Management (BLM) manages over 10 percent of the United States (270 million acres or 108 million hectares) which is 41 percent of the Federal Lands. Lands managed by BLM are known as Public Lands as opposed to the National Forest Lands which are Federal Land Policy and Management Act (FLPMA) of 1976 which in Sec. 102 states ??that it is the policy of the United States (7) ?that management be on the basis of multiple use and sustained yield??, the mission of the BLM is rapidly refocused under the concepts of biodiversity, ecosystems, and ecosystems-based management for the benefit of the environment, human needs and socio-economic values.

Jan 1, 1994

By C. A. Rowland

Introduction Ball mills are lined drums, either cylindrical in shape or modified cylinders that have either one or both ends of the shell, consisting of conical sections, that rotate about the horizontal axis. Fig. I I shows a cylindrical mill, Fig. 12 a conical ball mill, and Fig. 13 a Tricone ball mill (Hardinge tradename). Steel or iron grinding media, generally in the shape of spheres, are used to grind the ore to the specified product size. In order to obtain more contact area for grinding and to simulate the shape of worn balls, balls have been made with two concave surfaces diametrically opposite each other. Some concentra¬tors, such as Erie Mining Co., have used slugs cut from worn and broken rods to supplement the balls in ball mills and save money otherwise lost as rod scrap. Cylindrical and conical shapes have been tried instead of balls, but balls remain as the most common shape grinding media used in ball mills. Ball mills were a logical development from the earlier pebble mills that used hard natural pebbles such as flint pebbles or sized ore pebbles (obtained from the ore itself) as grinding media. In the early 1900s36 it was found that when cast iron or cast steel balls were used in place of flint or ore pebbles, the mills drew more power and gave greater production capacity. Advances in technology have resulted in the manufacture of ball mills up to 18 ft diam inside shell, drawing up to 8,000 hp. Ball mills are employed to grind ores, especially the more abrasive ores, to finer sizes than can be produced economically in other size¬reduction machines such as roll crushers, hammer mills, and impactors. Ores can be ground dry-dry grinding-or in a slurry-wet grinding-using ball mills. Dry grinding nominally refers to less than I %v moisture by weight. If the moisture content increases by several percent, dry grinding capacity is significantly reduced as shown in Table 17. The usual range of solids content in wet ball-mill slurries is from 65 to 80% by weight. Wet grinding is used to prepare the feed material for unit opera¬tions such as flotation, magnetic separation, gravity concentration, and leaching that require a slurry of liberated valuable mineral and unwanted gangue particles. Dry grinding" is employed to produce feed for agglomeration, pelletizing, and pyrometallurgy processes that require feed that is dry or nearly so and for finely ground industrial mineral products used in the dry state. Dry grinding is also used when minerals cannot be dewatered economically to the required moisture level or when the ground product reacts unfavorably with liquids. For example, cement clinker must be ground dry. Dry grinding requires about 30% more power than wet grinding for comparable size reduction .28 The total power required in a dry¬grinding ball-mill plant including drying may be double that required for a wet-grinding plant. Grinding-media and liner consumption in dry grinding reported as pounds of metal consumed per kilowatt-hour per ton of ore" is 10-20% of that used in wet grinding. The Wabush pellet plant, Point Noire, Que.3o reported ball consumption dropped from 6.3 lb per ton of ore ground to 2.5 lb per ton of ore ground when they converted from wet to dry grinding, and a 30% increase in power consumption. A number of comparisons made on wet and dry grinding of cement raw materials show metal consumption in dry grinding to be 10% of that in wet grinding. The capital costs for wet grinding are generally lower than for dry grinding. When thickening and filtering of the wet-ground product are required, dry grinding may have a lower capital cost. With open-circuit grinding the ball-mill discharge passes directly to the next processing step without being screened or classified and no fraction is returned to the ball mill (Fig. 14). In closed-circuit grinding the ground material, undersize, in the ball-mill discharge is removed either using a screen or a classifier with the oversize being returned to the mill for additional size reduction (Fig. 15). The over¬size material that is returned to the ball mill is called the circulating load. Open-circuit ball-mill grinding requires more power than closed¬-circuit grinding for products containing similar amounts of top-size material. The less the amount of oversize allowed in the product, the longer the ore must remain in the ball mill when grinding in open circuit. This increases the production of extreme fines and thus the consumption of more power. The power required for open-circuit ball-mill grinding can be estimated using the multipliers listed in Table 18 and knowing the power required for closed-circuit grinding to yield the desired product particle size. For example, assuming the desired grind size is 90% passing some specific top size, open-¬circuit grinding would require 1.40 times the power to achieve similar results as closed-circuit grinding.

Jan 1, 1985

By John F. Shea

When Dick Robbins asked me to give a keynote address, I felt it might be in order to review some of the practices and techniques used in tunneling and work up to some of the concepts now used to mine and construct tunnels. Needless to say, there have been many changes in tunneling technology and also many changes in the Safety Codes controling them. An example of this latter change was the fact that in 1931, my father excavated the Owyhee Tunnel No.5 with gas locomotives. He was able to do it because there was no natural gas in the rock and the ventilation system was designed to free the tunnel atmosphere from powder and exhaust fumes. Today OSHA rules would never permit gas locomotives underground, although there were no incidents on this particular project. One of my first tunnels, as Project Manager, was the Clear Creek Tunnel for the Bureau of Reclamation. Our method of handling muck was the Conway Muckers. This unit was a valuable improvement to the hand mucking of previous generations. The Conway was also used in our Navajo Project, in New Mexico. Unlike the Clear Creek project, where water was one of our main problems - up to 7,200 gallons per minute, at a pressure of 395 pounds per square inch - in the Navajo Tunnel, we had a mixed face which caused a major problem with our support system. In one of the Navajo Tunnels, there was a major cave-in, when the squeezing shale came in during an attempt to strengthen the support system. Although moving shale was not a constant problem, there were areas where the shale was five to six feet above the spring line.

Jan 1, 1976

By Daniel J. Turk, William R. Banning, John Paterson, Thomas J. Weitz, Michael L. Clark

Introduction Pegasus Gold's Montana Tunnels mine is a large surface gold, silver, lead, and zinc mine located 37 km (23 miles) south of Helena in southwestern Montana. Milling operations commenced in late March 1987. The concentrator achieved full production in May that year after undergo¬ing several modifications and improvements. Production for 1988 is projected to be 2.48 t ,80,000 oz) of gold, 34.2 t (1.1 million oz) of silver, along with 8.5 kt (9400 st) of lead, and 22.6 kt (25,000 st) of zinc from lead and zinc concentrates. Mill throughput currently averages 11.7 kt/d (12,900 stpd), compared to the origi¬nally designed capacity of 11.3 kt/d (12,500 stpd). Cash production costs, net of bypro¬duct revenues for the third quarter 1988, were below $3.20/g ($100 per oz). The introduction of ore to the concen¬trator marked the culmination of 12 years ex¬ploration, evaluation, and development. Initial milling operations encountered diffi¬culties with the circuit as designed. As a result, several modifications and improve¬ments were made to the plant. This article is intended to assist other companies developing projects of this type and complexity. It discusses the mine's de¬velopment from feasibility and construction through start-up. The operation, as it exists today, is reviewed including the modifica¬tions and improvements made since start-up. History The mine is situated in Jefferson County, MT within the historic Corbin-Wickes min¬ing district. Underground mining began in 1864 and flourished until the early 1900s. Between 1864 and 1900 the vein deposits in the district produced about $41 million dol¬lars worth of lead and silver ores containing subordinate amounts of zinc, copper, and gold. At various times, three smelters were operated in the area.

Jan 1, 1989

By Vernon Miller

In 1986, a subsidiary of Echo Bay Mines Ltd. purchased the McCoy property, 55 km (35 miles) southwest of Battle Mountain, NV. Echo Bay then launched a two-year, $1 15-mil lion capital program to build a mill to extract gold and silver from high-grade ore mined in two open pits - the original McCoy pit and a more recent discovery, the Cove. Engineering on a new mill began in 1987 with the facility coming on line in 1989. It was comprised of a semiautogenous (SAG) mill and ball mill grinding, leaching, flotation, counter current decantation (CCD) and Merrill-Crowe processes. The mill started up on McCoy ore, a rocky material that settled very well in the new CCD thickener system. Production levels were maintained at about 7.2 kt/d (8000 stpd). The situation changed as the McCoy ore played out and increasing amounts of material from the Cove pit-with its oxide, sulfide and carbonaceous ore - were fed to the mill. The problem was the high clay content of the Cove ore. The new ore immediately created production problems, mainly loss of control on the CCD thickeners. With the McCoy pit slated to be shut down in early 1991, it was essential to find a way to efficiently process the Cove ore. Initially, the only way to regain control of the thickeners was to slow down the processing rate and increase flocculant usage, thereby getting the solids to settle in the thickeners. But this had its cost. Production dropped to about 5.9 kt/d (6500 stpd), additional operators were needed and flocculant costs, on occasion, would reach an unacceptable $80,000 per month. Overfloccing led to viscous

Jan 1, 1993

By Gordon C. ‘Bud’ Presley

The pumice and pumicite review arrived too late for the June issue. Estimated domestic production of pumice and pumicite in 1992 was up compared with 1991, according to US Bureau of Mines statistics. The Bureau estimate of domestic production of 400 kt (440,000 st) for 1991 turned out to be 401 kt (442,000 st) of actual production, the lowest since 1989. In 1992, the estimated the value of pumice and pumicite sold or used in the United States was $11.6 million, up only $200,000 from the 1991 value. There was an increase in the mine, or mill, average price from the five-year low of $11.44/t ($10.37/st) in 1986 to $26.42/t ($23.96/st) in 1992. Estimated tonnage produced in 1992 was 439 kt (484,000 st). Apparent domestic consumption was 642 kt (707,000 st), up 135 kt (149,000 st) from 1991. Imports for consumption were estimated to be 215 kt (237,000 st) last year. Domestic production supplied 68% of the demand in 1992, with net imports supplying the remaining 32%. This compares with domestic production sup¬plying 53% and net imports supplying 47% of the apparent consumption in 1988. Exports were about 900 t/a (1000 stpy) for the period 1984 to 1988. In 1989, exports rose sevenfold to 7 kt (7700 st) and to 20 kt (22,000 st) in 1990. The Bureau of Mines reported that the amount of pumice and pumicite exported in 1992 was 12 kt (13,200 st). It is assumed that these exports were from the Pacific Coast because imports, mainly from Greece, continued to be a market factor on the East Coast and the Gulf Coast States in the United States.

Jan 1, 1993

By J. G. Dunn, A. S. Ibrado, D. J. Hollbach

Gold production is one the most important mineral process­ing industries in Western Australia. The gold is usually found in either weathered oxidised zones, or associated with sulfides. In the latter case, oxidation is achieved by some form of chemical processing, such as roasting or more recently bacterial leaching. In all cases, gold is dissolved out of the oxidised matrix by cyanide solution. Pyrrhotite is an iron sulfide generally represented by the formula Fe1_,S where O x 0.125. Three different structural forms ofpyrrhotite are known to occur naturally, stoichiometric FeS (troilite), hexagonal pyrrhotite Fe1_xS and monoclinic pyrrhotite Fe7S8 [I]. Substitution of iron atoms within the pyrrhotite structure by other ions of similar atomic radius such as nickel, copper, cobalt and manganese is also possible [2], and often observed in natural pyrrhotite samples. Pyrrhotite is abundant in Western Australian sulfide mineral deposits which are commonly associated with gold bearing ores. It is generally believed that pyrrhotite has a detrimental effect on gold recovery from such ores, as it is commonly claimed to decompose in alkaline, aerated cyanide solutions to form iron cyanide complexes, thus consuming free cyanide, and to form sulfur/oxygen containing species such as sulfate, through oxidation of sulfur in pyrrhotite, which thus consumes oxygen. From information gleaned from the literature, claims about the high reactivity of pyrrhotite in cyanide appear to be based on early cyanidation studies performed on gold ore samples which have been poorly characterised, and contained impurities such as pyrite, arsenopyrite and chalcopyrite, which may have also contributed to the levels of species observed in solution [3-5]. In addition, the investigations do not take into account that iron atoms in the pyrrhotite structure may have been partially substituted by other atoms that influence the solubility of pyrrhotite in cyanide solution.

Jan 1, 1995

By C. Richard Tinsley

Investment activity in the Pacific Rim of Fire has seen a phenomenal surge from epithermal gold exploration. Individual countries are attractive because of lower up-front minerals lease costs (United States, Canada), better geological prospects (Chile, Indonesia), or cheaper equities (Australia). Since the October 1987 stock market crash, large companies are shedding minority joint venture interests to focus on core businesses. Some vertical integration is evident in Japanese coal investments in the Pacific Rim. This article does not cover the aluminum industry oligopoly, which is multinational in character. Nor does it cover the many Canada to US minerals investments. It will be interesting to see if the new US-Canadian free trade treaty will reverse that investment flow. Cross-border investments The following examples should dispel any doubts about the significance of cross-border investment in the Pacific Rim. BHP, Australia's largest public company, has invested $US 4.5 billion in the JS for the following transactions 1984 $2 billion Utah Minerals 1984 $500 million Energy Resources 1985 $745 million Monsanto Oil 1987 $400 million 50.7% Hamilton Oil 1989 $378 million Pacific Resources Placer Dome, mainly through 76% owned Placer Pacific, has the following gold mine developments in hand including the date the project began, Placer Dome's ownership, project name, its location, and capital expenditures: • 1985, 53.1 %, Kidston, Queensland, $A 139 million; • 1989, 37.9%, Big Bell, Western Australia, $A135 million; • 1989, 60.6%, Misima, Papua New Guinea, $C252 million; • 1990,45.5%, Granny Smith, Western Australia, $A95 million; • 1991, 50%, La Coipa, Chile, $US276 million; and • 1991, 25.3%, Porgera, Papua New Guinea, $C1.1 billion. Australia's Western Mining Corp. (WMC), through a foray of off-market acquisitions in late 1987, acquired con¬trol of five Canadian gold companies for a $A488-million outlay ($C452 million) as follows:

Jan 1, 1990

By Richard Switalski, Thomas Scotese, James Morrison, Randall Essex

OVERVIEW Like many older metropolitan areas across the U.S., the Greater Cleveland area is served by a combination of separate sanitary sewers and combined sanitary-stormwater sewers. In 1966, a Master Plan was implemented by the Northeast Ohio Regional Sewer District (District) to minimize pollutant overflows into adjoining Lake Erie and its tributary rivers and streams, that would occur during wet weather events. One component of the Master Plan was to reduce the load on combined sewers by constructing new conveyances to intercept separate sanitary flows and convey them to upgraded regional waste water treatment plants for priority treatment. As part of a 15-year program to construct 74 km (46 miles) of new interceptors, the West Leg Interceptor (WLI) Project intercepts flows from four decommissioned waste water treatment plants (WWTPs), and conveys those flows to the new Southwest Interceptor (SWI), which conveys the flows to the Southerly Treatment Plant, a tertiary level treatment facility. The WLI Project was built in sections to permit decommissioning of the WWTPs as each segment of the interceptor system was completed and put into service. Flows from the first of the four WWTPs was converted to the WLI in November, 1992; the final connection was achieved in July 1994. The WLI comprises 9.1 km (5.7 miles) of 21 to 29 m (70 to 95 ft) deep mainline interceptor tunnel, four vortex drop structures, approximately 2 km (1.3 miles) of shallow tunneled connector sewer to the mainline, approximately 2.9 km (1.8 miles) of open cut connector sewers, five construction work shafts (not including one work shaft constructed under SWI Contract 6 at the north end of the WLI), and 15 access shafts.

Jan 1, 1995

By Tim Neil, O&apos

Copper price rebounds A worldwide scramble for diminishing supplies of copper has pushed the red metal's price to more than $2.20/kg ($1 per lb), the highest level since 1980. The first nine months of 1987, world copper production was down 1% from a year ago, while consumption rose 7%. Demand has been fueled by economic strength in several countries, particularly Japan, South Korea, and Taiwan. The price has also been helped by interest rate declines that have prompted manufacturers to build up inventories. In addition, China is expected to be a major buyer in 1988 as it stocks up for its five-year industrialization plan. There appears to be no sign of any significant increase in copper supplies until later 1988 or early 1989. This means supplies probably will remain tight for some time even if world economic activity slows. Strikes and other production problems slowed world output of copper in 1987. As a result, copper supplies are the lowest they have been in 14 years. World inventories currently total 625 kt (690,000 st). That is not much above the 585 kt (645,000 st) level reached in 1974 when the price of copper rose to a record $3.84/kg ($1.74 per lb). This makes all the more interesting what Chile, the world's largest copper producer, has been doing lately. Chile has been the world's largest copper producer since 1982, when it overtook the United States. Table 1 shows Chile's copper grades, reserves, and production at its mines. Chile contains roughly one-fifth of the world's copper reserves. At present production rates, these reserves would last for more than 100 years. In 1986, Chile produced 1.4 Mt (1.5 million st) of fine copper, up 3%. Chile accounted for 17% of the world's copper output.

Jan 1, 1988

By William T. Folwell

INTRODUCTION - CURRENT RAISE BORING PRACTICES Raise boring is a comparatively new tool to the underground mining and construction industries. Underground raise boring, in this country, had its beginning in the early 1960's, in the iron mines of Michigan. These first raises were bored 40 in. and 48 in. in diameter with maximum lengths of approximately 200 ft. Since the late "sixties", the raise boring machine has been accepted by the mining industry as an important machine to increase excavation rates, provide adequate ventilation airways, improve safety, and reduce costs, to name a few of the benefits. The construction industry is now accepting raise boring techniques for hydro-electric power projects, water supply projects, tunnel projects, waste water projects, etc. Raise drills are now capable of boring raises to diameters as much as 12 ft, in the softer rock types. Raise lengths range from 200 ft to as long as 1700 ft. Bored raises, in general, are bored at dip angles ranging from 45º to 90º, depending on the requirement of the raise. The general practice is to drill the pilot hole down to an opening below and up ream the pilot hole to the required diameter. The pilot hole (9-7/8", 9", 11", or 12-1/4" diameter) is generally drilled with compressed air to remove the cuttings. In the event that ground water is encountered and removal of cuttings is impeded, a foaming agent is added to assist in lifting the cuttings. When excessive water is encountered, water is used as the circulation medium and the cuttings are removed by pumping. During the reaming cycle, the cuttings fall by gravity and are removed from the bottom of the raise by conventional methods.

Jan 1, 1997

By C. W. Nelson, A. G. Moody

This paper describes the design, equipment selection and onshore fabrication of the largest facilities module ever installed in a single lift in the Gulf of Mexico. Measuring 35.5 x 70.1 x 21.3 m (110 x 230 x 70 ft), the three operating level structure houses Frasch process sulfur mining equipment and weighed nearly 49 million N (5500 tons) during lift. An operating weight of over 74 million N (8300 st) combines with bridge reactions acting on the in-place module and results in a total operating vertical load on the supporting platform deck of about 107 million N (12,000 tons). A Vierendeel truss structural system was selected for framing the module to allow maximum equipment access and routing flexibility for complex, insulated large diameter process piping. Research and finite element analysis of crimp plate siding behavior in shear was carried out during selection of siding type for the module. An interactive design sequence was carried out that incorporated in-place operational criteria, onshore fabrication considerations, complex process equipment requirements, plus load-out, ocean transport and offshore lift analyses. Project schedule dictated that design, procurement and fabrication be fast-tracked with parallel activities accelerated to the maximum extent possible. Design and construction of this largest and most complex module was done concurrently with 52 other components of the project including jackets, decks, process and living quarters modules as well as two 122 x 22 m (400 x 72 ft) sulfur tankers. Total tonnage of steel structures for the project exceeded 90.7 kt (100,000 st). The 5 kt (5500 st) mining module proved the viability and costeffectiveness of larger modules than commonly used in Gulf of Mexico waters. This technical advance can be matched with economic opportunity for other large-scale offshore installations.

Jan 1, 1994