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By R. H. Hanewald

The processing of EAF baghouse dusts from both the specialty steel industry and low alloy producers will be reviewed. In addition, the INMETCO process has been modified so as to recover nickel, chromium and iron from many other waste materials such as pickling solutions or filter cakes, spent nickel and chromium plating baths, carbon refractories from pickling tanks, and super alloy wastes including baghouse bags. INMETCO is the only North American company presently capable of recycling nickel-cadmium batteries by high temperature metals recovery. This, along with the details of INMETCO's BDAT Test Program and results performed to pre-agreed standards with the EPA will be discussed.

Jan 1, 1992

By E. J. Ostrowski

The history of ironmaking dates back to 4000 BC when an open forge was used to produce a sponge iron from a heap of ore and charcoal. Even at this early date, man recognized the need for a fuel and a reducing agent. As time passed, man developed new methods of producing iron; however, these developments were centered on the mechanical aspects, and as a result, a furnace evolved which is now known as a blast furnace. During this period of evolution,' man struggled with the only fuel he knew - charcoal. Some 5600 years passed before any attempt was made to use coal as a fuel for the blast furnace. The use of coal had its beginning in 1619 when Dud Dudley of England first attempted its use in the blast furnace; as a-fuel and as a reducing agent. Coal charged to the top of the blast furnace was not satisfactory. The volatile matter of the coal choked the flues, and also the poor, strength qualities would not support the weight of the ore. Breaking up of the coal compacted the stock column to an impervious mass. One hundred and sixteen. years later in 1735, Abraham Darby of England introduced coke as the fuel and reducing agent in the blast furnace. The acceptance of coke as a fuel for the blast furnace was not spontaneous. As late as 1870, one-half of the pig iron product of the United States was made with anthracite coal, 30 per cent with raw bituminous coal and coke, and 20 per cent with charcoal. However, by 1910, 96.2 per cent of the blast furnace product was made with coke, 2.4 per cent with anthracite and coke admixture, and 1.4 per cent with charcoal. Coke became the standard fuel in the blast furnace. As the art of ironmaking progressed, more pig iron was produced and more coke was consumed. Coke oven capacity was progressively increased. In 1910, there existed approximately 4,000 coke ovens; in 1960 there was approximately 15,000 ovens. This increase in coke oven capacity was necessary to feed the mammoth appetite of modern blast furnaces.

Jan 1, 1962

By Clifford W. Schultz

Four native copper ores, two conglomerate and two amygdaloidal, were tested to determine their amenability to sorting. One of them, a conglomerate ore from the Kingston mine, showed a high sortability as evidenced by an 80 pct recovery of copper in the minus 6-, plus 1/2-inch size range. Using the Kingston as a case example, the cost of mining and concentrating was estimated for a standard case and five optional cases involving varying degrees of sorting. This analysis indi¬cates that the production cost per pound of copper can be reduced on the order of 1 to 1-1/2 cents per pound along with a significant decrease in capital investment.

Jan 1, 1975

By T. Gonzales

Magma Metals Company operates an Outokumpu Flash Furnace, four Peirce-Smith Converters and two Anode Casting Wheels supported by six Anode Vessels. Process gasses from the Flash Furnace and Converters are treated in a two train/double contact sulfuric acid plant. Sulfur capture for the smelter is +98%. The new flash furnace was commissioned on July 7, 1988 and has processed over three million tons of sulfide concentrate without a major rebuild. Magma has gained unique experience in the operation of the flash furnace at high oxygen enrichments with low off-gas dust loadings. New procedures have been developed to maintain secondary material generation at equilibrium and for maintaining metallurgical control of operating parameters. This paper describes all the procedures, operation control parameters and operating philosophy that has allowed Magma to achieve superior performance.

Jan 1, 1992

By Carrie Smith

This fall, the Society for Mining, Metallurgy & Exploration (SME) will begin a new chapter by welcoming its new executive director and chief executive officer (CEO). Melissa Russell, a successful nonprofit executive with more than 30 years of experience in global association management, officially stepped into the role on Sept. 29, 2025. She succeeds longtime CEO Dave Kanagy, CAE, who is retiring in March 2026 after more than two decades of transformative leadership. “I am honored to begin this new chapter with SME and inspired by the vital role this community plays in shaping our world,” said Russell. “I look forward to listening, learning and working alongside this remarkable group of professionals to advance innovation, champion sustainability and create a brilliant future together.” Russell brings a deep portfolio of leadership experience in technical and industrial associations to SME. Most recently, she served as executive director of the Institute of Electrical and Electronics Engineers (IEEE) Computer Society, the largest of the IEEE’s 39 societies, with more than 50,000 members across 150 countries. During her tenure, she strengthened global engagement, advanced financial sustainability and launched strategic humanitarian initiatives.

Oct 1, 2025

By Hung-Kei H. Lam

Rubidium and cesium were discovered by Bunsen and Kirchhoff about 100 years ago. These two elements are by no means as abundant as sodium or potassium. Their occurrence in minerals is widespread, but usually as minor constituents. There is no known specific ore containing rubidium, but rubidium is found replacing the. other alkali metals in .some minerals.. Perhaps the highest concentration of rubidium found in one mineral is in lepidolite, KLi(OH,F)2A12Si3010, which contains from 0 to 3.5%*Rb20. 0n the average, lepidolite contains about 0.5% Rb20. Other sources of rubidium are carnallite, KMgCl3?6H20; pollucite, H20.2Cs20.2AI203.9SiO2; rhodizite, R20.2A1203-3B203; biotite, H2K(Mg,Fe)3A1(Si04)3; sea water, rivers, marine and fresh water organisms, soils, and plants. Cesium, even though it is the least abundant. of the stable alkalies, has a specific ore. Pollucites a hydrous cesium aluminum silicate, may con¬tain as high as 42.5% Cs20, although again, the average Cs20 content is about 25% among the handful of known deposits. Other cesium-containing minerals are lepidolite and carnallite. Rubidium and cesium are very much like potassium in chemical properties, consequently, they are hard to separate. Because of the difficulty in obtaining pure cesium and rubidium compounds coupled with the scarcity-of supply, prices were high and the two elements enjoyed little industrial use. Just before the Second World War, the world. supply of rubidium and cesium, from Germany, was a matter of tens of kilograms per year and prices were about

Jan 1, 1959

By N. H. Coates

The Bureau of Mines designed, constructed, and operated an 18-inch- diameter, fluidized-bed combustor for highly caking coals to evaluate the method for possible application to power generation. In initial tests, combustion was carried on successfully, with carbon losses as low as 0.8 percent. Coal ash was not satisfactory as a bed material, but mullite proved to be acceptable. Overall heat-transfer coefficients from the bed (-20 + 50 mesh mullite) to a water-cooled tube were about 75 Btu/hr-ft2- °F.

Jan 1, 1970

By Dan Nilsson

Rock from underground mining has a size-distribution, that varies from mine to mine, depends on rock conditions and mining method and on how much explosives etc, that is used. Because at the high costs, associated with full-scale studies of the size-distribution of blasted rock in a mine, detailed information is seldom published about such size-distributions. However, some facts can be noticed. Rock from tunnelling and other types of development work has a fine fragmentation. Cut-and- fill mining is a method that normally results in blasted rock with an advantageous size distribution. Block caving on the other hand often results in large amounts of boulders, since blasting of the rock is normally not used.

Jan 1, 1985

By J. A. Matar

Mineralization features at the Bajo La Alumbrera porphyry copper-gold deposit are discussed as they affect gold leaching and copper flotation behavior. Alteration-mineralization geological blocks are assigned in the zoned body, and each required its own geostatistical model. They extend horizontally outward from potassic to the phyllic-propylitic boundary, vertically from surface down to 300 meters, and include oxide, secondary enriched, and hypogene mineralization. Copper and gold values are calculated geostatistically from assays and all available geologic-mineralogic information. Reserves are classified for each geologic block and for each open-pit bench into leach or float ore according to metal grades, geologic characteristics, and supergene-hypogene relations. Finally ore extraction behaviour is briefly described, grouping gold and copper reserves into metallurgical families.

Jan 1, 1986

By D. Kidd, T. Tucker, L. M. Indergard, S. Davis

"INTRODUCTION The Lisbon Valley Mining Co LLC (LVMC) operates the Lisbon Valley Mine located in SE Utah. The mine includes 3 open pits, 150-acre heap leach pad, and solvent extraction electro winning plant (SX/EW). Approximately 25MM tons of ore has been mined since 2005. The ore is stacked on the leach pad to an average height of 55 feet and occurs as a blend of crushed material (minus 2”) and run-of-mine (minus 24”). The host rocks include Dakota and Burro Canyon Sandstones. Copper occurs finely disseminated in this material at an average grade of 0.46%. Minerology is 63% sulfide, 27% oxide, and 10% insoluble. Total inventory mined through 2016 is approximately 230 MM lbs. METALLURGICAL CONSTRAINTS Copper production is constrained by areas of low permeability ore and geochemical issues, including acid consumption. The intrinsic permeability the ore ranges from 10-2 to 1 Darcy. This permeability does not permit sufficient aeration by natural convection, therefore oxygen transport is largely limited to the mass of oxygen that can remain dissolved in leachates. As a result, greater than 50MM lbs of copper remains in inventory and occurs secondary sulfides, including chalcocite, covellite, and bornite. BIOLEACHING FUNDAMENTALS Due to the higher activation energies of secondary sulfides, notably covellite and bornite, a strong oxidant is needed to remove an electron from the metal-sulfide lattice and release the copper cation into the sulfate solution. This process occurs most efficiently using the strong oxidant ferric iron (Fe3+). Ferrous iron can be oxidized to ferric iron by oxygen via chemical oxidation or biologically; which was measured to occur 500,000 times faster when using microbes versus oxygen.1 The microbes that oxidize ferrous iron (Fe2+) to ferric iron are naturally abundant and will develop biomass when appropriate conditions and nutrients present themselves. Most nutrients are provided by the ore/irrigation (P,N,Mg, Ca & Fe2+) while the others (O2 and CO2) are provided by air. In most cases, oxygen transfer is the rate limiting step during copper sulfide leaching. 1 Lacey & Lawson 1970 FORCED AERATION SYSTEM DESIGN This paper describes LVMC’s efforts to bioleach the remaining inventory by forced aeration. Two basic aeration systems were installed and tested including vertical wells and horizontal piping. Vertical System The vertical system was installed using the mine’s blast hole drilling rig. Wells were constructed using 4” high-density polyethylene (HDPE) connected to 4” polyvinyl chloride (PVC) screens similar to conventional bioventing wells. Ore samples were collected at each well location to characterize the spatial distribution of the remaining inventory along with fines (minus 200 sieve), pH (acid consumption), and iron. A block model of each analyte was developed and mapped in 3D."

Jan 1, 2017

By Tina Živec, Žibertm Marko

"Civil design has changed in many ways using BIM (Building Information Modelling) technology in its most recent form. Such a technological breakthrough has not been successful so far in traditional oriented Engineering Geology, except in the Mining and oil exploration markets with budgets on much different scales. The fact is that paper, pencil and hand drawn sketches are still state of the art and a perfectly acceptable way of producing subjective geological interpretations.Using the technology available, we have developed a set of procedures and techniques to combine the latest surveying and remote sensing technology, GIS based tools to manage large data sets and BIM modelling techniques in a 3D geological environment.To demonstrate the concept, the forthcoming Karavanke tunnel project (Slovenia) has been chosen for testing techniques. The project, with its rich documentation from previous phases and new data available, represented a good basis for testing the possibility of digitizing geological data while communicating with Leapfrog Geo 3D modelling algorithms. This type of procedure would then allow engineers at the construction site to use the 3D model for prognosis and evaluation, while upgrading it daily with new information.INTRODUCTIONThe Karavanke tunnel crosses the mountain range on the border between Slovenia and Austria. As the geology is very complex due to its imbricate fan structure, a series of geological – geotechnical and hydrogeological investigations have been performed since the 1970s, resulting in extensive project documentation. The construction of the 7864 m long single-tube two-lane tunnel took place in the late 1980s. Currently, the tunnel represents a critical point on Pan-European Transport Corridor X (Salzburg, Austria – Thessaloniki, Greece), with insufficient safety measures. The need for improved traffic flow and enhanced road safety led the Slovenian government in 2013 to a decision to continue with activities regarding the construction of the second tunnel tube 70 m east of the existing tunnel.The extensive project documentation from previous investigation phases and construction of a new tunnel tube presented a good base for a pilot project, in order to test the emerging new technologies for data acquisition, processing and analysis in a 3D environment. Several products are available on the market for data acquisition and 3D geological modelling, but only a few can be applied in the civil design, as well. Developing a standard format and procedure for implementing the 3D geological model using new remote sensing techniques, 3D geological modelling tools, such as Leapfrog Geo and BIM (Building Information Modelling) modelling techniques would move the Engineering Geology, Geotechnics and Tunneling to a more advanced level. 3D modelling tools would therefore not only be used as a visualization tool, but would also contribute to in a more efficient communication and workflow between the professionals involved in the project."

Jan 1, 2016

By T. Murt, C. Oñate

In recent years, environmental considerations have played a prominent role in underground construction projects. The management and disposal of the muck produced by an EPB TBM is an important aspect of this topic. Soil conditioners with higher biodegradability and lower eco-toxicity have been specifically developed to comply with strict regulations in several countries around the world, with the goal of reducing the environmental impact of tunneling on soil, water, and living organisms. This paper highlights innovative sustainable soil conditioners and how they differ from traditional products. Additionally, environmental considerations for TBM projects around the world and in North America will be examined.

Dec 1, 2022

By Paul H. Muller

The configuration of remote seal DP cells for specific gravity measurement is described and the calculations for calibration are presented. The error induced by ambient temperature change is investigated as it relates to capillary dynamics and electronic considerations. Instrument specifications are discussed in the light of accuracy requirements and the calculation of error vs. ambient temperature is described.

Jan 1, 1984

By S. L. Bessinger

Introduction In the design and specification of a ventilation system for an underground mine, it is necessary to make reasonably accurate estimates of the pressure losses in the various airways of the mine. These estimates can be made with little difficulty for open airways with simple geometric cross-sections, such as those cut by continuous miners or tunnel-boring machines. The situation is much different on a longwall face, where the airway's complex geometric cross section and the presence in the airway of obstructing equipment having a variety of shapes make it difficult, if not impossible, to estimate pressure loss using traditional methods of calculation. Head losses in mine entries are calculated using Atkinson's Equation. [22H= KPLQ (English) H= KP 3Q (SO (1) 5.2AA] where H = pressure loss, in. of H2O (Pa); K = friction factor, lbf•min2/ft4 (kg/m3); P = perimeter, ft (m); L = airway length, ft (m); Q = airflow quantity, ft3/min (m3/sec); and A = flow cross-sectional area, ft2 (m2) In this equation, the friction factor, K, is an empirical constant that describes the aerodynamic roughness of the airway. Typically, the K-factor for a given airway is determined by measuring the factors H, P, L, Q and A in Equation (1) and calculating K. Tables of friction factors calculated in this way are found in textbooks and handbooks that deal with mine ventilation analysis. Unfortunately, very few K-factors have been measured on longwall faces, and the accuracies of those that have been measured are entirely site specific, because of the wide variety of equipment found on longwalls. The development of a technique for prediction without mine-site measurements of the friction factor for any longwall face, regardless of its configuration, will thus be very useful in the design of ventilation systems for mines in which longwall mining is practiced. Calculation of pressure losses using Atkinson's Equation (1) and empirically determined K-factors provides accurate and useful approximations in cases where the airways have relatively simple cross sections. However, a careful analysis using the principles of fluid mechanics shows that such calculations are based on two assumptions that are not strictly correct when there are obstructions in the airway. The first assumption is that the air velocity distribution in the cross section, particularly around the perimeter, is uniform. This assumption results from the fact that the tabulated K-factor values found in the literature are based on field measurements with uniform conditions. Such uniformity does not exist in longwall airflows. The second common assumption is that the K-factor, and corresponding head loss, is independent of the Reynolds Number (NR) for a given flow. In fact, this assumption is not strictly correct, and is particularly erroneous in the case of irregular protuberances into the airflow, such as those found on a longwall face. The errors arising from the assumptions may be avoided by using K-factors calculated by a newly devised method, described below, which takes into account the fundamental principles of aerodynamic drag analysis. This new technique has two advantages: first, it is flexible enough to model any longwall, regardless of equipment configuration; second, it employs terminology and equations familiar to those who perform mine ventilation analysis, using K-factors, for which ventilation engineers have an intuitive understanding, rather than drag coefficients. To provide guidance for development of a longwall drag model, data were taken on two modern longwalls operating in substantially different conditions. Pressure measurements at Mine B were made with 200-foot (61-m) sections of 1/8-in. (3-mm) diameter plastic tubing, attached to a Dwyer Magnehelic gauge. Pressure drops were measured in 200-foot increments down the face, and summed to give the drop for the entire face length. This technique was found to produce small, repeated errors because of the number of segments required to span the longwall. At Mine A this problem was avoided by using a single, continuous, plastic tube for the entire face length. The psychrometric properties of the air were measured for both Mines A and B. A calibrated, standard vane-anemometer was used to measure the airflow on both faces. Finally, numerous dimensions were measured on both faces, and face profile drawings were obtained to allow detailed evaluation of the face equipment geometry. From this information, accurate evaluations of the average wetted perimeter and average area of the longwall face airways were made. Since the airflow is not confined to inside the powered supports at all points along the face, a quadratically weighted average of the airflows measured at various stations along the face was calculated: [n2Qavg =Qi Ii / It(2)i=1] where [Q, avg = average airflow for analytical purposes, ft3/min (m3/sec); Q= airflow at station i, ft3/min (m3/sec); 1= length of segment represented by Q, ft (m); h= length of longwall face, ft (m): and n= number of quantity measurement stations.] The quadratic weighting scheme was chosen because the

Jan 1, 1992

By Kenneth S. Johnson

Large reserves of high-purity Permian gypsum crop out and are at shallow depth in three regions of western Oklahoma, Gypsum beds in the Blaine Formation are 10 to 30 feet thick and are generally 95 to 99 percent pure, whereas the massive Cloud Chief Gypsum is 20 to 100 feet thick and is 92 to 97 percent pure. Reserves are estimated at 48 billion short tons. Eight companies in the State currently mine about 1 million tons of gypsum each year for wall- board, for plaster, as a retarder in portland cement, and as a soil conditioner. Future uses, under more favorable market conditions, may be in the chemical industries, largely because of gypsum's potential as a source of sulfur. Three thick sequences of Permian rock salt underlie most of western Oklahoma at depths of 30 to 3,000 feet. Individual salt beds 5 to 30 feet thick are interbedded with thinner layers of shale and anhydrite. Five large salt plains, fed by natural springs of saturated brine constitute additional resources of salt. Reserves, estimated at 20 trillion short tons, are virtually untapped, and, ironically, about 90 percent of the salt used in Oklahoma is imported. Chemical data on the rock salt are lacking, but evaporated salt produced from brines is generally 97 to 99 percent pure. Current production from three solar-evaporation plants is used mainly for stockfeed, recharging water softeners, and deicing roads. Potential uses are in the chemical industries, where salt or brine is a necessary raw material.

Jan 1, 1973

By J. O. Down

St. Marys Cement Company operates two cement plants in Southern Ontario, Canada. The original plant, located at St. Marys, a town 20 miles north of London, Ontario, began operations in 1912. A second plant, located at Bowmanville, Ontario, 40 miles north east of Toronto, began operations in 1967. In June 1973, St. Marys Cement began feasibility studies of modernizing their wet process plant at St. Marys. Over its 61 year life, this plant had been expanded from 335 m. tons per day (375 s. tons per day) to 2000 tpd (2200 s. tons per day), with all the original equipment retired several years ago. The objectives of the feasibility study were to develop a conceptual plan for a dry process plant and to de-, fine the appropriate process method and equipment. This study proceded along four fundamental fronts: a) How could the clay raw material, a glacial till with up to 20% boulder content, be processed in dry form and beneficiated? b) Were the raw materials suitable for the most efficient form of dry process kiln, the 4-stage preheater? If a 4-stage preheater kiln is selected should it utilize a precalciner? c) What type of clinker cooler should be used? d) What environmental constraints would be imposed on this plant and how can they be reconciled with capital and operating limitations? During the feasibility stage, pilot plant work was done on two possible solutions to the clay problem, operating plants in Europe were visited and discussions held with kiln and mill vendors.

Jan 1, 1977

By Peter Schmäh

Pipe-jacked tunnels are considered for a large variety of applications from water and sewage, cable ducts to challenging pipeline crossings or landfall construction. In most cases, Slurry MTBMs are used to carry out these projects. Continuous further development of technological features for pipe jacking plays a key role to overcome feasibility limits and to meet increased project requirements. As tightened safety regulations in terms of accessibility and escape routes increasingly limit the options of small segment lining tunnels in the diameter range below 4 m, 13 feet, pipe jacking has partially taken over. During the last decades, a growing number of large-diameter and long-distance pipe jacking projects have been completed. Experienced contractors successfully use the technological capacity of the MTBM equipment. The machine and cutting wheel have to be designed according to the ground conditions. Bentonite lubrication and interjacking stations are key factors to keep push forces moderate. For long pipe jacking drives, a so-called push module can be installed in the machine can, to be able to switch from pipe jacking to segment lining in case the machine gets stuck. The overall aim is to provide clients and the tunnel construction industry with safe and economical solu- tions for successful project execution. This presentation will present international milestone projects and will also discuss the latest developments in automation, related opportunities and limitations, and the future of digitization and artificial intelligence in pipe jacking.

Jun 23, 2024

By Paul M. Jr. Musgrove, Thomas D. Henderson, Mark B. Nesbitt

Introduction History. Vat leaching operations began in October 1953. Since that time, about 4.2 million tpy of oxide ore averaging 0.72% copper have been treated. Economic conditions, improved operating methods, and increased production facilities have allowed the ore grade to be reduced over the years. The average ore grade was 0.84% Cu during the 1950s, 0.68% in the 1960s, and was 0.50%, or below, during the 1970s. By 1961 some 35 million tons of oxide ore had been removed from the mine, exposing the underlying sulfide ore body. A 5,000¬ tpd concentrator was constructed and put into operation to treat the sulfide ore in November 1961.1 The original crushing plants were used for both oxide and sulfide ores, but by adding a third crushing shift, six days per week, the total ore treated was raised from 12,000¬17,000 tpd. Dwindling oxide ore reserves prompted expansion of the concen¬trator to 13,500 tpd in 19672 so that exhaustion of known reserves of both ores would occur at nearly the same time. Again, no additional crushing facilities were added during the expansion. It was anticipated that the treatment rate of oxide ore would be reduced from 12,000 to about 9,000 tpd. Instead, the addition of two crushing shifts per week and improved operating procedures have increased the treatment rate to 26,500 tpd-13,000 tpd to the leaching plant and 13,500 tpd of sulfide ore treated in the concentrator. Geology. The Yerington ore body is a disseminated porphyry copper deposit where a quartz monzonite porphyry intrusive contains the major part of the ore values, with some copper mineralization found in the adjacent granodiorite and other rocks. The geology of the Yerington Mine was aptly described by J. R. Wilson (Mining Congress Journal, June 1963). The ore body is one of few porphyries where geologic conditions were optimum for the formation of an important concentration of oxidized ore with minimum economic effect of secondary enrichment. The principal mineral in the oxidized zone is chrysocolla, which irreg¬ularly occurs dispersed throughout the rock and as narrow seams along fissures. The grade of the ore in the oxidized zone is generally the same as in the primary sulfide below, although some massive concentrations do occur as vein-like fissures and as filling of breccia zones. Cuprite, tenorite, melaconite, malachite, and azurite occur but are not abundant. Lying between the chrysocolla horizon and the primary sulfide is a transition zone whose limits are irregular but seldom with vertical range exceeding 50 ft. Chalcocite, cuprite, melaconite, native copper, and chrysocolla occur superimposed on the primary sulfides. Chalcocite replacement of chalcopyrite, as well as secondary chalcocite, do occur. The primary sulfide minerals are chalcopyrite and pyrite, with minor amounts of bornite and covellite. There are no appreciable gold or silver values, and molybdenite has rarely been noted.

Jan 1, 1985

By Donald E. Steif

The Tower Mill is a vertical stirred mill using steel balls or pebble grinding media. This device is designed for fine grinding at significant power savings when compared to conventional tumbling mills. This paper describes the machine, its operation, application, and performance for wet grinding.

Jan 1, 1984

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

6.1 PRINCIPLES OF PLANNING GROUTING OPERATIONS FOR SHAFTS, DRIFTS AND TUNNELS The planning of integrated grouting of vertical, inclined or horizontal underground workings in saturated, fractured rock includes the following phases: 1. Planning for the collection and analysis of geological information concerning the characteristics of the struc¬tural deformation(s) that produced the details of the frac¬ture system being studied. 2. Planning for the acquisition of initial hydrogeological information produced by the flowmetric and hydrody¬namic investigations (described in Chapters 2 and 3). 3. Computing the initial engineering calculations of the ex¬pected or planned dimensions of the isolation curtains for each water-producing stratum (fracture zone) and the expected areal spreading pattern of the grout in each of these water-producing horizons. The selection of the lo¬cations, the initial optimal number of holes and the vol¬ume of grout needed to create each isolation curtain also are included in this phase. 4. The selection or development of a grout formula for the specific hydrogeological conditions that exist at the site. 5. Identification of a resource (deposit) required to produce the grouting clay. The planning of grouting operations must be supported by the acquisition of all necessary initial data, which in¬cludes: 1. The geographic location and administrative responsibil¬ity (ownership and regulatory constraints) of the mineral deposit and any peculiar features of the area where ex¬cavation of the underground workings is planned. The surface relief, the hydrographic network of the region, and the local climatic conditions must be evaluated prior to planning for buildings and waste discharges if any. 2. The economic status of the region, the available trans¬portation, the nature of the energy base, the water and heat supply sources, the location of construction mate¬rials suitable for making the grout all must be deter¬mined. 3. The geological and hydrogeological characterization of the work site must be accomplished. 4. The technical features of the shaft(s) and underground works must be characterized. 5. The organization's principal initial design plan features and the construction sequence of all buildings and struc¬tures must be developed. These must be correlated with the organization and schedule for excavating the under¬ground works. The technical characterization of a shaft or of other un¬derground workings must contain information about the purpose of the workings, the design depth, the planned excavation procedures along with cross-sections, and the parameters of the lining, if a lining is required. The geology of the section to be penetrated must include geological data obtained from existing exploration activities at the mineral deposit and conclusions about the shaft ex¬cavation conditions obtained from drilling technical moni¬toring holes and exploration drillholes. The interpreted stratigraphic regime, including the thickness, strike and dip of each stratum, must be shown in cross-section. The tec¬tonics, including all geologic structures, must be presented by cross-sections. The bedding orientations must be used to identify anticlines, synclines, monoclines and faults. The lithologic composition and the drillability of the rock com¬prising the section must be examined. The character of the rock's fracture pattern must be presented. Finally the per¬meability, porosity and storage characteristics of each aqui¬fer must be determined by in situ hydraulic property testing (injection tests or pumping tests, as described in Chapter 3). The fracture pattern of the indurated rock is presented in reports derived from geo-exploration borehole studies and from studies of outcrops on the surface and in all nearby underground workings. The extent of each fracture zone is mapped in three dimensions. The distribution of fracture apertures, the nature of the fractures according to each litho¬logic stratum, and the relationship of bedding planes to each fracture system are all determined. Acoustic logs, downhole cameras, caliper logs, oriented core, and any other appro¬priate geophysical tools available are used to construct the data base for this purpose. The hydrogeological characterization of the geologic section to be penetrated by the workings is based on the analysis of hydrogeological data collected during the drill¬ing and testing of technical monitoring holes and from in¬formation on ground water inflow collected during the ex¬cavation of nearby underground workings, if such workings exist. The hydrogeologic characterization must contain the following information: 1. The identification of all aquifers, their thicknesses and their locations in the section, 2. The distribution of hydrostatic head in the ground water in three dimensions, 3. The coefficient of saturated hydraulic conductivity (per¬ meability) of each aquifer, 4. The porosity and storativity of each aquifer, and 5. The projected rate of water inflow into the shaft, drift or tunnel for each aquifer prior to grouting. This step is used to identify the zones to be grouted.

Jan 1, 1993