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By Frank X. Bellini, Marc Saidman, Allen J. Hulshizer, Anthony J. Stewart

The cooling water system for the Seabrook Nuclear Station involves 10.25 Km (33,623 ft.) of 6.7 m (22 ft.) OD hard rock tunnels, over one third of which is under and terminating over 1.6 Km (a mile) out in the Atlantic Ocean. The major excavation is accomplished by two TBM's concurrently driving the intake and discharge tunnels which by mid-January, 1981, completed 92% of the excavation. This paper presents the case history of the excavation work and describes a computer program developed to evaluate major activities of TBM excavation on a shift-by-shift basis. Details and summaries of the production efforts are computer plotted in different forms to facilitate the evaluation of production performance, trouble spots and trends.

Jan 1, 1981

By D. Koren, D. Carson

Disposal of smelter slag may be subjected to regulatory leach tests to deter- mine whether they produce a leachate containing contaminants in excess of stated limits. Contaminants that are most often of concern for smelter slags are As, Cd, Cr, Cu, Ni, Pb, and Zn. The present regulatory tests are based on the assumption of storage in a municipal landfill dump, where wastes are in contact with organic acids. Since slags are not normally stored in such an environment, these tests do not adequately assess the true leachability potential of the slags. Ten slags from six major Canadian nonferrous sulfide smelters were studied. Leaching was done following regulatory and modified tests, including leaching in columns with synthetic acid rain for a 25-week period. The slags were characterized by mineralogical analysis before and after leaching with synthetic rain. The synthetic mineral equivalents (hereafter referred simply as phases) in the slow cooled slags consist of fayalite grains and dendrites, amorphous glass, some spinel and small quantities of complex sulfidic prills and metallics. The phases in the quenched slags consist of amorphous glass, some spinel and small quantities of complex sulfidic prills and metallics. Some metal dissolution took place during leaching. However, for all leaching tests, the concentrations were always below the regulated limits. In general, the content of the potentially leachable elements remains unchanged in the fayalite, in the amorphous glass and in the spinel, whereas the quantity of sulfidic prill and metallics decreases after leaching. Therefore, most of the deleterious element dissolution comes from the sulfidic prills and metallic phases. Metal dissolution increases by crushing the slag, probably due to an increased exposure of the sulfidic prills to the leaching solutions.

Jan 1, 1998

By Nobuharu Isago, Atsushi Kusaka, Kosuke Kawata

"Rock tunnels suffered severe damage from recent large earthquakes in Japan, despite the empirical knowledge that tunnels tolerate earthquakes better than surface structures. Seismic design for rock tunnels have therefore become a fascinating subject for many tunnel engineers. However, even basic behavior, including deformation mode and loading magnitude for hard-rock TBM tunnels and conventional ones during earthquake, has not been fully understood due to lack of actual seismic data; while EPB and slurry TBM tunnels take account of seismic design in Japan if needed. This study conducted a dynamic measurement in an actual road tunnel constructed in rock. The strongest aftershock of “the 2011 off the Pacific coast of Tohoku Earthquake” (Mw = 7.1) occurred during the measurement, providing precious data to examine deformation mode of the tunnel. Considering the acquired data, numerical analysis was also performed to calculate loading magnitude to cause fatal fracture of tunnel lining during earthquake.INTRODUCTIONLarge earthquakes can pose serious risks even for rock tunnels, as evidenced by large earthquakes in the past. In the large earthquake in Japan in 2004 (Mw = 6.6), some tunnels suffered severe damage including collapse of their permanent lining (Mashimo 2005), even though the tunnels were constructed in ordinary ground conditions except for fault, fracture zone, portal area, etc. To mitigate such risks, mechanical behavior of tunnel lining during earthquake should be clarified; however, even basic mechanical behavior including load-bearing capacity of the tunnel lining against earthquake is not fully understood.Numerous damage modes were actually observed due to various ground conditions (eg. Wang et al. 2001; Mashimo 2005). Among them, major tunnel damage caused by earthquakes can be grouped into three main modes as shown in Figure 1, excluding extremely unusual conditions such as a tunnel located on a stratum boundary where the stratum stiffness varies significantly. TYPE-I shows flexural failure or crack by bending moment at shoulders; TYPE-II indicates compression failure or flexural compression failure at crown; and TYPE-III illustrates compression failure at sidewall, mainly observed on tunnel that was constructed using steel ribs and lagging method. These damage modes, Type-I, II and III, are predominantly caused by shearing, horizontal compressing and vertical compressing deformations of the surrounding ground, respectively (Kusaka et al. 2011; Asakura et al. 2001)."

Jan 1, 2016

By Wilbur I. Duvall

This paper discusses general principles of underground opening design based on the concept that with the necessary input data regarding the geology of a site, the physical properties of the rocks, the in situ stress field in the rock and the general geometry of the underground openings, the rock formation can be classified into six major categories which are useful for design purposes. Using simple theories for the criteria of failure for rock In compression and tension, general elastic theory for stress concentration around openings and In structures, beam theory for stability of roof rock, and elementary inelastic theory for creep phenomena, basic design techniques are outlined for the various rock classifications. The need for field evaluation of the theoretical design is stressed and some general techniques that should be used in these field evaluations are discussed. Where appropriate, the need for additional studies and research is noted.

Jan 1, 1977

By Jerome C. Motz

Los tipos, tamaños y número de trituradores empleados en un sistema completo de reducción varían con factores tales como: el volumen de mineral que debe ser procesado, tamaño de la mina, tamaño máximo del todo-uno, dureza del mineral y de su roca soporte y tamano requerido para el producto final. Básicamente, sólo existen tres tipos distintos de equipos de trituración para la etapa primaria. Estos son: machacadoras de mandíbulas, trituradores giratorios y molinos de impactos, y cada uno tiene unas características operativas propias. Las trituradoras de mandíbulas operan por aplastamiento de las rocas entre dos paredes, una fija y otra móvil, en una cavidad determinada. Han sido ensayadas variaciones en la caída y movimiento, pero la mayoría de las máquinas del tipo Blake tienen un ángulo de trituración de aproximadamente 27° entre las mandíbulas fijas y móviles. El principio de trabajo del triturador giratorio es muy simple. Si un cono es montado en un eje vertical y la parte superior del eje se sitúa estacionaria mientras que la parte inferior gira excéntricamente, el cono tendrá también un movimiento excéntrico. Si el cono se coloca en una carcasa conveniente se movers aproximándose y retrocediendo de las paredes que lo contienen a la vez que gira. Si el cono y las paredes de la carcasa son suficientemente robustas y pesadas, cualquier cosa apresada entre ellas será triturada. Al lado de los incrementos de la eficacia desarrollados por las continuas acciones de trituración y el curvado de las caras de trituración, los trituradores del tipo giratorio tienen las siguientes ventajas: 1. La gran abertura de alimentación disponibles en comparación con trituradores de otro tipo. 2. El alto rango de tamaños y capacidades, que se encuentran entre 600 y 6.000 toneladas por hora. 3. Generalmente no es necesario control en la alimentación. Dumpers de 150 a 200 toneladas descargan directamente en las actuales instalaciones.

Jan 1, 1982

By D. A. Swanson, G. Danziger, G. Savci, C. Barnes

Hydrologic assessments of existing and proposed waste rock stockpiles are typically required for application of operating and closure permits at mine sites. A hydrologic assessment should include an evaluation of climatic data and prediction of the net infiltration rate through the waste rock stockpile surface. In addition, hydrologic factors such as the time required to deplete available moisture storage in the stockpiles and the time required for potential seepage to migrate to the underlying aquifer should also be predicted.

Jan 1, 2000

By Alfred C. Maevis

On October 27, 1972, ground was broken for the first section of the Second Avenue Subway, a two track underground rail system that was to extend from the south end of Manhattan at Broad Street, north to the Harlem River at 126th Street, a distance of eight miles. This ground breaking was the culmination of 67 years of planning for an east side subway in Manhattan. Following the successful opening of New York City's first subway on October 1904 there was a mass planning for additional subways throughout the City. In 1905 approval was requested of the Board of Estimate for a First Avenue route and Third Avenue route. The Third Avenue route was approved but when action was not taken to start construction, the approval in accordance with the conditions under which it was given was withdrawn. Further reports and plans for an east side subway, usually to be constructed on Third Avenue, were issued in 1920, in 1922, in 1924, etc. Those of us who had worked on the Sixth and Eighth Avenue subways in the thirites and early forties left to join the armed services of World War II in the expectation that we would return and start work on designing and constructing that east side subway. The reason that much of the thrust in those years was to construct on Third Avenue was that there was a Third Avenue "EL" and business interests of the City were in favor of its demolition. This demolition was accomplished in 1955 so one of the incentives to construct on Third Avenue was removed; but an east side subway was always part of the master plan of the City.

Jan 1, 1976

By Henry R. Tiedemann, Eugene H. Skinner, George E. Wickham

The need for ground support research in underground excavation has been well identified in recent years by many studies in the subject area (1, 2, 3, 4, 5). A typical conclusion from these reports (1, 2) is that the drill-and-blast (D & B) mining or the mechanical tunnel boring machine (TBM) methods are expected to be continued in use for the needs of tunneling in the immediate future. The Organization for Economic Co-operation and Development (OECD) sponsored in June, 1970 advisory conferences on Hardrock and Soft Ground Tunneling (3,4). These two reports were based on the combined response received from approximately six hundred inquiries from 17 nations in the summer of 1969 and which concerned all aspects of present-day inadequacies of, and the needed improvements in the technology of tunneling. Some specific problem areas in ground support and tunnel linings were: 1) the ability to successfully predict rock loads that must be carried by the support system; and 2) inadequate methods for determining the load-carrying capacity of the various support systems. Therefore, the problem area is clearly defined by these two objectives: 1) To find means of predicting rock loads; and 2) To develop methodology for determining safe, efficient, and economical ground support design practices. Noting these research needs from a great number of sources, the United States Bureau of Mines, Spokane Mining Research Center (SMRC), acting as agent for the Advanced Research Projects Agency (ARPA), undertook contract research effort in these problem areas. Work by Jacobs Associates has studied a methodology for specification of temporary artificial support based on the dependent sequence of mining operations and prediction of various rock conditions in the pre-excavation stage. The major thesis is that rock behavior is the

Jan 1, 1997

By Leo J. Laughlin

In recent years, heavy mining applications have put an increasing demand on conveyor pulleys. This has resulted in the increased usage of turbine pulleys. The designs of these pulleys are often based upon the failures seen in the conventional welded designs, overlooking the interaction of pulley-shaft system components as the profile of the disc is modified. This paper will focus on the development of balanced design philosophy using case studies of turbine pulleys in heavy mining applications, evaluated with FEA.

Jan 1, 2002

By A. B. Seal, C. J. Bise

Excessive noise has long been a hazard in the coal-mining industry. Studies conducted during the mid-1970s and mid-1990s consistently show that hearing loss within the mining industry persists, in spite of regulatory requirements and sampling technology advances. When the U.S. Mine Safety and Health Administration's new health standards to protect miners from hearing loss took effect in September 2000, the potential developed to reevaluate sampling approaches with regard to noise-source hazards, such as task-based methods. This paper describes results from a Pennsylvania State University research project now being conducted at an underground coal mine. Each occupation was reviewed, and a list of tasks or processes performed by each was generated. Traditional personal noise dosimetry, static sound pressure levels and equipment sound mappings have been performed. Job tasks indicating the greatest levels of daily noise exposure and contributions to noise dose have been identified and will be discussed.

Jan 1, 2003

By A. C. Lyons

Foreword Sir William Halcrow and Partners are responsible for the design of the Ahmed Hamdi Tunnel as Sub-contractor to the Contractor, Osmac, a joint venture formed by the Arab Contractors, Osman Ahmed Osman and Company of Cairo, and Tarmac International of London. Throughout the paper the term "Client" refers to the Employer, the Ministry of Housing and Reconstruction, which latterly became the Ministry of Development and New Communities of the Arab Republic of Egypt. At the time of writing of this paper only a few hundred metres of the tunnel have been driven, but the design is substantially complete. Introduction The Ahmed Hamdi Tunnel, which is at present under construction, will be a two-lane tunnel crossing beneath the Suez Canal approximately 15 km north of the town of Suez at El Shallufa. It will become the first permanent connection across the Canal since 1967 and is visualised by the Egyptian Government as an early step in the development of the Sinai Desert along the eastern bank. The tunnel will carry pipelines for treated water and electric power cables to assist in this development. The tunnel, which will be 1640 m in length, is being driven through hard clay and mudstone using a tunnel shield fitted with hydraulic excavators. The primary lining is being constructed with precast, bolted reinforced concrete segments with an internal diameter of 10.4 m. The road deck structure and the walls of the air supply ducts beneath it are formed with precast, reinforced concrete units erected immediately behind the shield.

Jan 1, 1979

By Moussa Wone, William Levy, Christopher Caruso, Zachary Spera, Thomas Hennings, Richard Stacy Kinchen, PE Joel Kantola

"Owners constructing new combined sewer overflow (CSO) systems increasingly seek to have project design lives of 100 or more years in response to the growing size, complexity, and cost to construct and operate these systems. However, a codified and universally accepted approach to achieving this goal in the United States does not currently exist since service life prediction is a relatively new technical field. This paper discusses how the 100-year design life requirement was met for the design of tunnels, drop shafts, and diversion structures in the DC Clean Rivers Program (DCCRP). The authors identify some commonly referenced publications, present the testing framework used by the program manager to characterize the CSOs, review the design criteria used to guide the 100-year design life development, and establish guidelines to be considered by program managers and designers when implementing similar extended design life requirements in other CSO control programs.INTRODUCTIONLike many older cities in the United States, the sewer system in the District of Columbia consists of both combined sewers and separate sanitary sewers. When the capacity of a combined sewer is exceeded during storm events, the excess flow, which is a mixture of sewage and storm water runoff, is discharged to the District’s receiving water bodies (Anacostia River, Potomac River, and Rock Creek) and tributaries through outfall structures. This excess discharge during storm events is called a combined sewer overflow or CSO. A total of 53 CSO outfalls for the combined sewer system are listed in DC Water’s National Pollutant Discharge Elimination System (NPDES) Permit issued by the U.S. Environmental Protection Agency (EPA). The DC Clean Rivers Program (DCCRP) will reduce the number of CSO outfalls into the Potomac and Anacostia Rivers and decrease flooding in DC neighborhoods during extreme rain events by capturing, storing, and conveying diverted CSOs to the Blue Plains Advanced Wastewater Treatment Plant."

Jan 1, 2016

By R. J. Puchala, B. M. Hawrylewicz

High voltage electrical discharge in water causes instantaneous release of accumulated energy into water. The experiment described in this paper regards the utilization of the electric discharge phenomenon in rock splitting. The pressure in the shock waves created by the discharge inside a small volume of water enclosed in solid rock causes stresses which can result in breakage of rock. Holes drilled in tested samples were notched in order to weaken the tensile strength of the rock and achieve a uniformly directed split. The influence of the notches was theoretically analysed and compared with experimental data. Results of the tests have been statistically elaborated in order to correlate the newly-created surface to the energy supplied.

Jan 1, 1986

By I. H. Keith

INTRODUCTION The following papers describe the extraction of nickel in the form of ferronickel from the Dominican laterites. The Falconbridge process was developed initially in the Falconbridge Metallurgical Laboratories and subsequently in several successive stages of pilot plant operation in the Dominican Republic. In the course of this work we came to believe that there is advantage in piloting on site. Mining, metallurgical and geo-political considerations all have major impact and that belief has been borne out by commercial plant experience. Lateritic nickel ores differ substantially, even within the same orebody, in their amenability to extraction methods. The Falconbridge process takes account of the heterogeneous nature of the deposit. Three years of geological field exploration led to ten years of process development after it was realised that the then known metallurgical processes would not give economical extraction from the Dominican ores. In addition to establishing the viability of a novel pyrometallurgical approach, it was necessary to verify design parameters for a very considerable scale-up to prototype commercial operation. The plant was eventually constructed within the original estimates of time and cost. Unique features contributing to the technical and commercial success of the project are described. Additionally, some bold extrapolation based on pilot plant work, eventually proved to be quite conservative. The mine is not a conventional open pit operation, but is perhaps better characterised as an earth moving job. Novel application of mobile equipment led to great flexibility in the mine, where operation of several mining fronts is essential to maintaining uniformity of ore feed to the plant. The practical aspects of almost total acceptance of the oversize run-of-mine material through a high-capacity unit,

Jan 1, 1979

By Safdar A. Gill

The technique of installing small diameter caissons (drilled piers) through soft clays and water bearing ganular soil is well established for constructing building foundations. Over the past few years, the method has been utilized for constructing a large number of shafts extending into rock for the Chicago's Deep Tunnel Project. This paper will discuss design considerations for obtaining the most economic ground support system, the impact of construction procedures on the support and lateral pressures for design. Construction procedures will be described along with performance of the shafts.

Jan 1, 1981

By G. Kemp Williams, Peter H. Dohms, Barry A. Hillman, Laurie A. Williams

Introduction In 1969, the US devalued the dollar. This action, over the ensuing four years, resulted in the deregulation of the price of gold. It also resulted in the rebirth of the US gold mining industry. The extreme inflation of the 1970s led to a dramatic increase in the price of precious metals. Consequently an increased interest in the exploration for and mining of gold has occurred. The enthusiasm began among the few existing producers who had survived on marginal profits during a prolonged period of fixed prices and constantly rising costs. Then came the newcomers. They saw the opportunity to enter an industry that historically had been characterized by small operations with low capital costs. It was not until the late 1970s that large mining companies recognized that investment in gold projects could generate cash flows that previously were associated only with world class deposits. Today, gold mining has tremendous appeal for the entire mining community, despite the dramatic price drop from 1980-81 levels. There are even major firms with no mining experience attempting to gain a foothold. But how easy is it to enter the US gold mining industry? What are the basic factors that influence the economics, and what are they likely to be for the remainder of the decade? How has the industry changed, and what is the composition of the active companies? An examination of the industry will help to answer these questions. US Gold Mining History The current gold rush is actually the fifth in the 208-year history of the US. The first recorded commercial mining of gold occurred in the Southeast in 1799. It produced from small, yet relatively rich placer deposits. The famous California gold rush of 1848 placed the US as a major producer with 62.2 t (2 million oz) of output by 1850. More than 93.3 t (3 million oz) was mined by 1853, equivalent to 63% of world production. But the richest placers were soon depleted. By the end of the Civil War, annual production had declined to well below 62.2 t (2 million oz). The government then relaxed prices and the industry again began to recover. The 1870-80s saw the great mines of the California Mother Lode and Grass Valley brought into production. This was followed by development of Nevada's Comstock, which later became one of the world's largest producers. Discovery of gold at Cripple Creek in 1892, coupled with technological advances in the use of large dredges in the California placers, resulted in significant production increases. And, with the expansion of the Homestake mine at Lead, SD, in 1898, US output again surpassed the 93.3-t (3 million-oz) level. By 1905, produc¬tion climbed to 124.4 t (4 million oz) with the addition of the Alaskan deposits, new output at Goldfields, NV, and the introduction of the cyanidation process. During World War I, the government forced the shutdown of gold mines. By 1920, total production had dropped to little more than 62.2 t (2 million oz). Many mines remained closed until the Great Depression. This, combined with the devaluation of the US dollar in 1934, resulted in a gold price increase to about $1.13/g ($35 per oz). Then began the fourth gold rush. By 1937, output had again surpassed 124.4 t (4 million oz). This was due to an increase in primary production and new byproduct production from base metal operations. Unlike the previous period of growth, however, this period was short-lived. The imposition of Executive Order L-208 in 1942 once again forced the shutdown of most mines. At the end of World War II, cumulative US gold production had

Jan 12, 1984

By Harold G. Arthur

INTRODUCTION Stillwater Tunnel construction is scheduled to start in the spring of 1973. The tunnel will be the site of an extensive program in rapid excavation research. This paper presents a progress report regarding our plans and relates the tentative decisions which have been made on how to carry out the research effort. The details are subject to revision as more input is obtained from other Government agencies interested in tunnel construction and from experts in industry and educational institutions. The objective of the research program is to bring about faster, less costly tunnel construction. This objective will be pursued through the development, improvement, and evaluation of various construction systems and investigative techniques. The need for this kind of program was recognized by the National Research Council, in its authoritative report "Rapid Excavation - Significance, Needs, Opportunities." (1) The Council advocated a federally funded $200 million, 10-year program to help meet the $69 billion market it expects in underground excavation between 1970 and 1990. The significance of tunnel construction to Reclamation is made clear by statistics of its tunneling experience and outlook. Reclamation has constructed 174 water conveyance tunnels totaling over 190 miles, Six tunnels are presently under construction which have a combined length of about 27 miles. There are on the books nine authorized, but unfunded tunnel projects totaling another 70 miles in length. Many more miles of tunnels are included in programs in various stages of investigation.

Jan 1, 1997

By A. J. P. Durham

Two low water flux test covers were placed on acid generating waste rack dumps at the Kidston Gold Mine in Queensland, Australia The climate is semi arid with distinct wet/dry seasons. The present study was initiated to develop a cover system to restrict infiltration. The design of the cover required optimization of water storage and evapotranspiration. Instrumentation was installed on both dumps to measure infiltration, moisture conditions and climatic parameters. Field measurements over the last three years indicate that both covers are performing satisfactorily. Selection of the final cover for closure is described on the basis of these results.

Jan 1, 2000

By James R. Arnold, John O. Marsden

Introduction Metallurgical accounting practices have been revolutionized over the past 15 years by the introduction of low cost personal computers and easy- to-use spreadsheet software. The Chimney Creek metallurgical accounting system takes full advantage of these advances. Data is provided by a PLC-based process control system, IBM data logging, operator logsheets and assay results from an automatatic plant sampling system. The data is organized and stored in an IBM PS/2 system using Lotus 123 and customized accounting programs to provide a comprehensive information management system. Metallurgical accounting for gold plants should ideally provide information for the following: a) Gold production accounting b) Day-to-day process control and operation c) Metallurgical evaluation and process optimization The Chimney Creek system provides all of these while overcoming specific accounting problems presented by SAG milling a coarse feed, grinding in cyanide and dump leaching run-of-mine material. Process description Gold Fields' Chimney Creek mine is located 47 miles northeast of Winnemucca, Nevada, at the northern end of the Osgood Mountains. Ore is processed by a combination of dump leaching and milling processes. The dump leach operation treats approximately 2.0 million tons per year of low grade ore (0.030 oz/ton). Run-of-mine ore is top dumped in 40 foot high pads. The use of three solution ponds for leaching enables an intermediate solution to be recirculated to the leach pad to reduce the pregnant solution flowrate and increase gold tenor of solution to the carbon-in-column (CIC) recovery circuit. Gold recovery is approximately 65% by this method. The milling process employs a SAG mill-ball mill grinding circuit to treat 900,000 tons per year at an average grade of 0.19 oz/ton. The ore is crushed to minus 12" in a jaw crusher prior to feeding the SAG mill. Cyanide is added to the grinding circuit to start gold dissolution as early as possible. After grinding, leaching of the ore is continued in a unique two-stage counter current decantation circuit 1 with 12 hours of leaching between stages . 85% of the soluble gold is recovered from the first stage thickener overflow in carbon columns (CIC). The second stage underflow is treated in a scavenger carbon-in-pulp (CIP) plant. Ultimate gold recovery is over 95%. Loaded carbon from both the dump leaching and milling processes is pressure stripped by hot cyanide solution and gold is recovered from the resulting solution by zinc precipitation. The precipitate is retorted for mercury removal and smelted into bullion. The plant flowsheet appears in Figure 1. Data collection Data for daily metallurgical accounting is obtained from three sources: a) Plant sampling b) Process control system c) Operator logsheets All of the data collection for metallurgical accounting is based around the 12-hour shift operating schedule. Sampling All critical sampling for gold production accounting is performed by automatic samplers (Figure 2). Heath & Sherwood two-stage sampling systems are used for sampling all slurry streams (cyclone overflow, CIP feed and CIP tails). The automatic sampler consists of a pneumatically operated primary sample cutter and a continuous rotary secondary splitter. The cyclone overflow and CIP tails are the most critical samples since these are used to calculate gold into and out of the plant on a shiftly basis. The CIP feed sample gives additional information concerning intermediate dissolution and recovery

Jan 1, 1990

By S. C. Bandis, N. Barton

During drilling of deep boreholes, local mechanical failures of the wellbore walls may occur due to the inability of the incompetent overburden and reservoir rocks to withstand drilling-induced changes of the stress environment. At present, borehole failure mechanisms in three-dimensional stress fields are poorly understood. Our experimental investigations based on 3-D physical models demonstrate the stabilizing effect of the intermediate principal stress. The latter may play a dominant role in maintaining the integrity of the stress relieving failure zone around the borehole. However, this will depend on the stress anisotropy, the angle of hole deviation and on the angle of internal friction. Our tests also indicate that the failure zone can accomodate high extensional strains. Finally, relations between the physical models and closed-form elasto-plastic solutions are discussed.

Jan 1, 1986