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By J. J. Krizek, A. H. Ditto

INTRODUCTION The Carr Fork property is situated in the Bingham Mining District located on the east slope of the Oquirrh Mountain range in Utah. The mine is adjacent to and lies west of Kennecott's Bingham Pit some 25 miles southwest of Salt Lake City. The Bingham Mining District contains one of the largest copper reserves on North America. Mineralization is associated with the Bingham stock which intruded through some 16,000 feet of sedimentary formations. Mineralization and a1teration zoning was developed radially from the Bingham stock with lower grade disseminated copper mineralization centrally located in the stock and adjoining quartzites. In addition, several garnetized limestone beds abutting the stock were also well mineralized. These skarns will be Carr Fork's initial mining target. Of the several limestone beds, only two, the Yampa and the Highland Boy have significant commercial interest at present. These beds which are 50 feet to 150 feet thick and are separated by 200 to 400 feet of quartzite were highly folded, thrusted and faulted, resulting in a dip variance from 20" to vertical. The predominant dip, however, is between 20" and 45" from horizontal. Quartzites surrounding the limestone beds were subjected to the same degree of tectonic activity with resultant intense fracturing and stress relief. The limestone beds which have a higher elastic modulus and which also experienced fracture healing during mineralization appear to have greater competency and strength than the quartzites. Four shafts nominally 3500 feet deep will provide access to the mine which will be a highly mechanized trackless operation, designed to produce 14,000 tons per hoisting day. First production is scheduled for the last quarter of 1979.

Jan 1, 1977

Overview In 1992, US coal supply and demand were shaped more by weather conditions and the export market than by the economic recovery. Preliminary data compiled by the Energy Information Administration (EIA) (Table I) show that coal production declined slightly to 902 Mt (994 million st). Consumption, though, rose by about 1 % to 811 Mt (894 million st), trailing the 2.1 %growth in Gross Domestic Product. Contributing to the decrease were a sizable drop in coal exports and some coal stock withdrawals. Consumption increased largely because electric utilities turned to coal to offset a shortfall in hydroelectric generation caused by a drought in the Northwest. Apart from this, the demand for coal-fired generation was dampened by a mild winter in the northwestern and northcentral United States, and a cool summer in the central and eastern parts of the Nation. Coal use by other consumers was about the same as in 1991. The average delivered prices of coal to all consuming sectors decreased slightly in 1992, reflecting soft coal markets (Fig. 1). US coal exporters experienced a set-back in 1992 mostly because of losses in the European market. Total coal exports fell by 6% to 93 Mt (103 million st), the first decline since 1987. Coal stockpiles at the end of 1992 were about 1.8 Mt (2 million st) below 1991, despite stock buildups by some electric utilities in the East. Stocks were built up as a hedge against possible coal miners' strikes after the Feb. 1, 1993, expiration of the contract between the United Mine Workers of America and the Bituminous Coal Operators' Association.

Jan 1, 1993

By R. Venugopal, T. C. Rao, M. Vanangamudi

Introduction The kinetic approach, to understand a process in detail, has been extensively used in the unit operations of many engineering systems. In green pelletization, the kinetic studies have been mainly oriented towards quantifying the growth mechanism (Kapur and Fuerstenau, 1964; Sastry and Fuerstenau, 1977; Capes and Danckwerts, 1965; Kapur and Fuerstenau, 1966; Kapur, 1972) and in a few cases, relating the rate of growth with added liquid content (Kapur and Fuerstenau, 1964; Newitt and Conway-Jones, 1958). Studies on the kinetics of green pelletization, in terms of the order of the kinetics are virtually non-existent. Sastry and Fuerstenau (1970) classified the agglomeration process into "free in-space" (dilute particle system) and "restricted-in¬space" (highly dispersed particle system) and noted that the former is a second-order process while the latter corresponds to a pseudo first-order rate process. The present paper attempts to describe the kinetics of green pelletization, in terms of order of the kinetics, using pyritiferrous shale as the material system. Experimental Pelletization experiments were carried out using -100 mesh pyritiferrous shale, in a laboratory disc pelletizer, at four intervals of pelletization time (15, 20, 25 and 30 min). The green pellets obtained were size analyzed. The size distribution of pellets has generally been represented by a single parameter such as the mean granule diameter, ratio of maximum and minimum pellet diameter, etc. In the present study, the d50 (50% passing size) of pellets has been considered to represent the size distribution of pellet. Kinetic equations The process of size enlargement is considered analogous to chemical reaction and a general kinetic equation of nth order for growth of pellets is given as

Jan 1, 1990

By Frank E. Dalton

BACKGROUND OF TARP The first settlement in Chicago in the 1830's was only 0.61 m (2 ft) above the level of Lake Michigan. The area was flat and swampy with local streams and rivers generally flowing to Lake Michigan. Repeated outbreaks of cholera and typhoid fever in the middle and late 1800's led to the reversal of the river system (away from Lake Michigan to the Mississippi River). As a result of this reversal, the Chicagoland area has had to restrict water withdrawals from Lake Michigan and has had to develop and implement water management programs in compliance with orders set down by the Supreme Court of the United States. In order to comply with restricted lake diversion requirements and to compensate for a natural setting which was essentially hostile to human habitation, it has been necessary to embark on the construction of massive public works which include the world's largest water filtration plants, the world's largest wastewater treatment plants and finally the construction of TARP. The rivers and streams in the Chicago area are the headwaters of the Illinois waterway system. These headwaters are heavily polluted from combined sewer overflows of the area. Present pollution control laws require that combined sewer overflows be eliminated so that the rivers and streams can be used by the citizens of Cook, Du Page and Will Counties and other downstream communities. Fifty-three of the older most densely populated communities including the City of Chicago, covering an area of 971 sq km (375 sq miles), are served by combined sewers. Combined sewers carry both raw sewage and stormwater. Because porous ground surfaces have been covered 1815

Jan 1, 1979

By Stephen J. Klein, Fernando Orduz, Randall J. Essex

INTRODUCTION The Keystone Tunnel project consists of widening an existing 7.3 m-wide (24 ft) highway tunnel to a width of 13.7 m (45 ft) to provide a three-lane, two-directional highway tunnel. The 13.7 m-wide tunnel was excavated in a staged drill-and-blast sequence through foliated schist, and was supported using pattern rock reinforcement and shotcrete. Notable project features are: the low vertical cover over the tunnel is less than the span of the opening; the project is the second instance where the Federal Highway Administration (FHWA) has accepted rock reinforcement as a permanent support element; and the project is the first instance where the FHWA is accepting shotcrete as the final interior surface of the tunnel. This paper covers the design and construction phases of the project. PROJECT DESCRIPTION The South Dakota Department of Transportation (SDDOT) retained Woodward-Clyde Consultants (WCC) to evaluate the feasibility, prepare construction plans and specifications, and provide construction management services for the widening of the Keystone Tunnel. The tunnel is located in the Black Hills National Forest area of South Dakota near the southwest corner of the state. The Keystone Tunnel is on U.S. Highway 16A, about 0.8 km (1/2 mile) north of Keystone which is approximately 32 km (20 miles) south of Rapid City (Figure 1). Highway 16A serves as an important transportation link between Rapid City and the Mount Rushmore National Monument which is located about 2 miles west of Keystone. The road through the tunnel consisted of two lanes about 7.3 m (24 ft) wide, and is being widened [ ]

Jan 1, 1989

By F. W. Brackebusch

Paste backfill is a pumpable, flowable, non-Newtonian fluid consisting generally of mine tailings and cement. Paste is prepared from dilute slurries of tailings. This is done by dewatering the slurries with conventional thickening or filtering. Backfill with a paste consistency is prepared by mixing dewatered tailings with cement and water to obtain the desired consistency of medium slump concrete. The paste is pumped with concrete pumps over relatively long (1 km or 0.6 mile) horizontal distances. More commonly, though, it flows by gravity down a mine shaft and then horizontally to the void being filled. Alternatively, cement can be added near the end of the pipeline. A certain proportion of fine particles <20 µm (<625 mesh) must be present in pastes but particles of aggregate can be added without significantly affecting the flow characteristics. Detailed testwork is necessary to design a system to prepare and place paste backfill. Testwork must determine dewatering characteristics, strength of mixes and friction in pipelines. Plants must be well engineered and controlled with PLCs because high quality control is necessary to prevent plugged pipelines and excessive pressures. Relatively small amounts of cement (3% to 5%) produce stiff backfill material 1.5 to 3.5 MPa (217 to 507 psi). The advantages of paste backfill in a mining operation include reduced mine dewatering, labor and materials savings, reduced tailings impoundment requirements, good support properties and reduced spills underground. In certain cases where more than 30% of the weight of the ore is sold as a product, the entire tailings stream can be placed underground. Paste technology can he used for surface or underground placement of tailings.

Jan 1, 1995

By R. Kuhns

The western coast of South Africa hosts a significant diamond resource within gravels associated with paleo-shorelines, channels and transgressive lag deposits. Diamonds were transported from the Kaapvaal Craton to the coast largely during the late Cretaceous via the paleo-Orange river systems. On shore and near shore environments are underlain by Proterozoic metamorphic and granitic rocks which provide gully and pot hole-type traps associated with paleo-channels, basins, bays and beaches. Inner and mid-shelf environments are covered by diamondiferous, poorly lithified upper Cretaceous and Tertiary sediments. The upper Cretaceous gravels were reworked along the sediment-starved coast during much of the Tertiary. Reworking was mainly achieved during marine flooding events which produced lag gravels along the paleo-shorelines. This process reworked the diamonds from low-grade Cretaceous gravels and concentrated them in higher-grade Tertiary gravels.

Jan 1, 1995

By Dapeng JIANG, Haixiao YANG, Shanxi WU, Jinwang Zhang, Guoliang XU, Yipeng ZHANG

The subsidence and deformation characteristics of strata with thick unconsolidated overlying strata is quite different from that with consolidated overlying strata. In this paper, based on the engineering background of large mining height mining in No.21 coal seam of Zhaogu No.2 Coal Mine in Henan Province, China, the PFC 2D numerical simulation and surface subsidence field observation was conducted to study the subsidence characteristics of strata with thick unconsolidated overlying strata (680m). Numerical model with dimensions of 2000m 750m was established and mining process with different working face lengths (150 m,200 m,250 m,300 m) and different working face advancing distances (40 m,80 m,120 m,160 m,200 m,240 m,280 m,320 m,360 m,400 m,440 m,480 m) were simulated. The surface subsidence characteristics and displacement field variation of rock strata under different working face advancing distances were observed and analyzed. The results show that with the advance of working face, the structure of subsidence-like basin will be formed in each rock stratum, and the influence range of mining on rock stratum is “bowl-shaped”, and the influence increases with the increase of advancing distance. The unconsolidated layer has a buffer effect on the mining-induced subsidence, and the buffer effect is realized by the expansion of the subsidence area. The movement law of strata under different working face lengths is similar to that under general mining, and the structure of subsidence-like basins will also be formed in each stratum. The influence range of strata after mining is ' bowl-shaped ', and the influence extent increases with the increasing of working face length; the internal subsidence of the unconsolidated layer is more intense and complex than that of the rock layer; the surface subsidence movement shows an obvious 'basin' phenomenon. The field observation data of surface subsidence also confirm the above results.

By I. WILKOMIRSKY, J. BECKER, F. Parada

CORRECTION

Jan 18, 2023

By D. R. Barker

The economic performance and competitiveness of a concentrator are dependent on the production of accurate real-time assay information and the manner in which it is utilized. Our paper describes the relative merits of on-and in-stream analysis techniques and the economic significance of Assay-Based control and Target Oriented Display Systems. At Outokumpu&apos;s pyhasalmi Concentrator, a Flotation Expert Display System aids the process operator to make better decisions and provides guidance on the most efficient method of achieving the desired control objectives.

Jan 1, 1989

By R. S. Dennen

A ground penetrating synthetic pulse radar system has been used by the Bureau of Mines to probe into a coal pillar to locate hidden structures or anomalies. Two types of velocity tomographic reconstruction methods, direct matrix inversion and iterative, were used to obtain a velocity image of the interior of the pillar. The reconstructed image showed the existence and location of a relatively low velocity material within the pillar. Core samples were taken from the pillar in a region surrounding the low velocity material. Analysis of the cores showed the presence of a clay vein coinciding with the low velocity material imaged in the tomogram.

Jan 1, 1989

By P. W. Hynes

Tg Soda Ash has operated a shortwall mining system in the Green River, Wyoming trona district since August. 1982. The system has proven to be highly productive. In an effort to reduce the relatively expensive development mining and Improve mine planning flexibility, a method to mine the return air tailgate during the shortwall retreat was tested beginning In April, 1987. As of June. 1988 four shortwall panels had been completed utilizing the advancing tailgate, producing over 815,000 tons.

Jan 1, 1989

By M. D. Dennis

Precious metal mineralization at Mt. Hamilton" Nevada is associated with a poly-metallic, garnet-pyroxene skarn system developed in middle to upper Cambrian calcareous shales, limestones and dolomites. The skarn is developed around the Cretaceous Seligman and Monte cristo calc-alkaline stocks. Gold is associated with intense retrograde alteration. Later epithermal quartz veins containing gold, silver, base-metal sulfides, and sulfosalts cross-cut skarn alteration. Westmont Mining Inc. is currently developing two mineralized zones containing a geologic resource of approximately 7.7 M.T., which averages 0.05 OPT gold and 0.5 OPT silver. The project will have an estimated mine life of seven years, with a planned annual production rate of 35,000 ounces of gold and 300,000 ounces of silver.

Jan 1, 1989

By J. W. Rains

Arizona Portland Cement Company&apos;s limestone deposit (called Twin Peaks) is approximately 4 miles southeast of the cement plant. The pant is adjacent to both the Southern Pacific Railroad and Interstate 10, about 17 miles northwest of Tucson. Placer claims were filed on the Twin Peaks deposits in 1923. From that time until the present, geologic data have been accumulated and evaluated to define more closely the quantity and quality of the limestones. The cement plant at Rillito was originally constructed as a one-kiln plant in 1949. Capacity was increased in 1952 and 1956, bringing the plant to an annual capacity of 44,000 tons of cement. In 1972 anther expansion program was completed, bringing annual cement capacity to 1.1 million tons. Geologically, the Twin Peaks contain formations ranging from the Precambrian Pinal Schist to the Pennsylvanian Naco Formation. Within this sequence, other exposed formations include the Cambrian Bolsa Quartzite and Abrigo Formation, Devonian Martin Limestone, Mississippian Escabrosa Limestone, and Pennsylvanian Naco Formation. Generally, the Escabrosa, Naco, and Martin Formation are the primary sources of limestone for manufacturing cement. The Martin Formation is basically a dolomite with bands of limestone, quartzite, and siltstone. These limestones are blended with outside source materials (clay and iron ore) to make a raw mix that will make specification cements. Several drilling programs were initiated between June 1946 and March 1982. Information derived from these programs provided the means to develop cross sections, geologic blocks, and grade blocks. The configurations of the geologic and grade blocks are the same. Block boundaries delineate the various formations on any particular level. Cursory inspection of the structure would indicate a broad syncline elevated on the west side. A more detailed examination shows that the structure has been complicated by several periods of faulting and intrusion of igneous rocks.

Jan 1, 1987

By Ben Ainsworth

I am going to discuss gold mining in Mexico during the last twenty centuries and try to suggest some of the futures that may develop in this industry in the next decade. Before I start this rapid review of such a long developing mining tradition, T must give credit to my sources. I an, talking here as an associate of Commoditi,c,s Reseflrch Unit Ltd. of London (better known as CRU) and as a principal of Ainsworth-Jenkins Holdings, an international mineral resource consultant group that has been exploring and deveJoping mining properties in Mexico and Central America for the last six years. CRU is the leading consultancy providing industry, governments and financial institu­tions with analysis and advice on the metals and the mining industries. They recently completed a three volume study entitled Gold Mining Opportunities in Mexico. This .<1tudy describes Mexico, the business environment, its geology and details its gold mining potential with property descriptions and maps. I have adopted and adapted the s11mmari.zed Mexican government statistics and rlrawn on some of the historical comments in these volumes. However, I do not have ti me (nor would it be appropriate) to be giving the details of the CRU work. Jorge Ordonez, whose shoes I am trying to fill today, was a major contributor to the study and it was an honour to be asked by Willem Beunderman, the Principal Editor, to step in when Jorge was unfortunately unable to be here. Jorge has been an enormous help to Ainsworth­Jenkins and a consistent friend of foreign investment in Mexico. Our own work in Mexico has allowed us to obtain a few useful in­sights into this most interesting slice 0f Mexican history.

Jan 1, 1992

By G. S. Walker, W. van Bronswijk

The accurate determination of free cyanide and metal cyanide complexes is of increasing interest to Carbon-in-Pulp (CIP) gold extraction operations as cyanide consumption is an important economic factor in the processing of complex ores. Characteristically complex ores are high cyanide consumers and these ores usually contain a range of sulfide minerals e.g. covellite (CuS), digenite (Cu1,S), chalcocite (Cu, S), Pyrrhotite (Feci-x)S), marcasite (FeS2), arsenic and antimony sulfides and zinc sulfide. These minerals can consume significant quantities of cyanide, e.g. various copper(!) cyanide complexes such as CuCN, Cu(CN)2·, Cu(CN)/ and Cu(CN)/ can form, depending on the cyanide to copper ratio in solution. Measurement of free and complexed cyanide is thus a prerequisite for efficient cyanide dosage control. Free cyanide can be determined potentiometrically or amperometrically but the methods are prone to error and are laborious [1]. Metal cyanide complexes can be determined by reverse phase HPLC, however, the species ofa single cation (e.g. Cu(CN)i-, Cu(CN);2 and Cu(CN)/) elute as one species [2]. Additionally the high amount of suspended matter in CIP leach solutions tends to introduce errors. Raman spectroscopy has the potential to be used as an in­line or on-line method of analysis. Free cyanide and the cyanides of gold, silver, copper and iron have been studied in aqueous alkaline solutions with a view to developing such a method. Spectra were obtained with a Bruker RFSl00 FT-Raman spectrometer using a 1064 nm laser delivering 800 mW of power at the sample. Depending on the solution, 250, 1000 or 4000 scans at 4 cm-1 resolution were accumulated. The cyanide stretches observed were relatively narrow (10 15 cm-1 full width at half height) and well resolved (Figure 1 ).

Jan 1, 1995

By Steven Hart, Graeme Hill

Boddington Gold Mine is situated 130km South-east of Perth, WA. The operation is managed by Worsley Alumina Pty Ltd. The Direct Leach plant was commissioned in mid 1986 at a rate of 3 million tonnes per annum. The plant was upgraded to 4.5 mtpa in 1988 and almost immediately to 6mtpa in 1989. Subsequent debottlenecking resulted in further increases in throughput up to 7mtpa by 1991. The SupergenelBasement Plant was constructed in 1990 and commissioned in early 1991. It was designed to process both copper-gold-silver anomalies within the laterite gold orebody and future hard rock ores. The Supergene plant incorporated control potential s­ulphidisation (CPS) and flotation ahead of CIL to remove a host of copper minerals. It operated in that configuration for 2 years at rates up to 350,000 tpa and average grades of 6.0 git gold and 1.2% copper. Gold recovery averaged 90.7% and copper recovery averaged 80.0% before the flotation circuit was decommissioned in March 1993. Quartz hosted high grade, free milling oxide and basement ores were discovered at the northern end of the orebody in 1990 and now comprise the major feed source to the basement plant. Oxide ores are currently milled at a rate of 550,000 tpa, while basement ores are milled at a 420,000 tpa rate. Metallurgical testwork during the feasibility stages indicated that up to 30% of gold in plant feed could be recovered by gravity separation methods. It was also highlighted that signifi­cant improvements in overall plant recovery and reduced operating costs would be achieved by maximising gravity gold recovery ahead of cyanidation. To achieve this, rec;iesign and upgrade of the gravity circuit was necessary.

Jan 1, 1995

By Satid Therdkiattikul, Donald A. Dahlstrom

Many gold ores are known to be refractory so that conventional cyanide processing will result in serious losses of gold and silver values. A significant percentage of these refractory ores contain aresenopyrite and the gold and silver is usually encapsulated within this matrix as well as the pyrite matrix. Normally the values can be recovered as flotation concentrates because of the sulfide compounds with recoveries of over 90% and at very good grades. In order to free the gold and silver, the sulfide matrix must be destroyed. This can presently be achieved by (1) pressure­temperature leaching (2) roasting and (3) bacterial leaching. All of them have advantages and disadvantages as well as an obviously increased cost for the each processing method. However, the increased gold and silver recovery should and must more than offset this cost. It is believed that gold recovery should be well over 90% from the concentrate so that overall recovery from the ore should be in the 90-95% region. All three processes must consider pollution control of both sulfur and arsenic to prevent contamination of air, water or land through improper waste disposal. This paper will concern only bacterial leaching and its necessary control of effluents and solids. Sulfur does not appear to be a problem as the element is oxidized to sulfate and is precipitated with an acceptable solubility. Other important improvements that must be controlled are arsenic and iron which are present as significant values (1-12 g/1 of arsenic, 20-40 g/1 of iron). In addition much smaller values of base metals, aluminum and magnesium can be present as well as other elements depending on the ore and these must be low enough after treatment to meet EPA regulations.

Jan 1, 1993

By Tom McDonald

Test work and plant operation confirmed that semiautogenous (SAG) milling is a viable option for Inco&apos;s milling rationalization project at Sudbury only if pebble crushing is included. Throughput increases more than 50% with pebble crushing incorporated into the comminution process. The finer product of high-speed pebble crushing provides further energy savings. The unconventional approach of pumping the pebble crusher product, while simplifying plant layout, has produced challenging wear problems. Inco Ltd.&apos;s Sudbury operations are located in northeastern Ontario, Canada, about 400 km (250 miles) north of Toronto. Inco has 11 operating mines around the rim of the Sudbury basin. The mines currently produce about 11 Mt/a (12 million stpy) of ore. This ore contains pentlandite, pyrrhotite and chalcopyrite in mafic host rocks. The milling facilities are designed to produce 108.8 kt/a (240 million lbs/ year) of nickel as concentrate. This concentrate is further processed through Inco&apos;s smelting and refining facilities to produce nickel, copper and several other metals. Before 1991, -200 mm (-8 in.) ore was shipped to two primary mills. The Clarabelle mill had a design capacity of 29 kt/d (32,000 stpd). Frood-Stobie had a capacity of 16.3 kt/d (18,000 stpd). The flowsheets of these mills were similar. Two stages of open-circuit crushing were followed by rod milling and closed-circuit ball milling. The product from the comminution circuit, at 12% +250 mm (+65 mesh), was first passed through magnetic separators to recover magnetic pyrrhotite. It was then sent through a four-stage flotation process. The first two stages produced a high-grade copper-nickel concentrate. The final two scavenger stages produced a low-grade copper-nickel concentrate containing nonmagnetic pyrrhotite.

Jan 1, 1992

By R. V. Hunter, J. F. T. Agapito

Introduction Significant increases in extraction ratio on the order of 6% to 8%, equivalent to as much as 142 t/m (48 st per ft) of panel, have occurred at the Big Island trona mine during the last five years with the help of a practical rock mechanics program. This increase in resource recovery also has contributed to improvements in productivity. Both conventional and continuous machine mining have been used to mine two flat-lying trona beds at depths of 250 m and 260 m (820 ft and 853 ft) in the Green River Formation in southwestern Wyoming. Conventional room-and-pillarmining originally was conducted in panels with extractions of 5 8% to 60%. The use of yield pillars has allowed the panel extraction in increase to 66%. Continuous machine mining was introduced, with a 64% extraction ratio using long, narrow pillars and wide rooms. Further improvements in resource recovery seem feasible in light of present mining experience. The rock mechanics program consisted of field instrumentation to determine the pillar and roof response hole locations. to mining and computer modeling to evaluate the stability of various layouts, which were then adopted for mining. Mining background Stauffer Chemical Co.&apos;s Big Island mine and refinery are located 40 km (25 miles) northwest of Green River, WY. Underground mining operations supply trona, a sodium mineral, to the refinery to produce soda ash. The room-and-pillar method is used to mine two 3 to 3.6 m (10 to 12 ft) thick, flat-laying seams at depths of about 250 to 260 m (820 to 850 ft). Mine production in 1986 was 2.3 Mt (2.5 million st). These were supplied through both conventional and continuous machine mining sec-

Jan 1, 1989