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By Peter Shuttleworth, Chris Fesq, Eddie Woods

Precast concrete segmental linings for tunnels have traditionally been designed with conventional steel bar reinforcement. In many cases, this has been satisfactory in service, but in others, major problems have occurred with the durability of the linings. Commonly this has resulted from corrosion of the reinforcement, mainly from chloridesin ground water, causing spalling of the concrete and leading to very expensive and disruptive repair programs. This paper discusses the relative merits of designing tunnel linings with steel fiber reinforcement to provide long-term durability, improved impact resistance and toughness for greater load capacity across joints, enhanced robustness during transportation and erection, plus improved fire resistance with the addition of polypropylene fibers. Guidance is also given in the selection of the correct fiber type, mix design and testing methods to establish the optimum dosage required and production techniques to ensure even distribution and compaction.

Jan 1, 2005

By R. D. Lord

This paper deals with the search for the best practical means of storing sulphide bearing tailings, where there is no residual excess of carbonate material. It is the common experience that the sulphide content decomposes, with the aid of bacterial action, and the resulting sulphuric acid escapes, along with any heavy metal solutes, through embankments that are usually porous to some degree. The problem is typified in the tailings of the uranium operations of Elliot Lake, Ontario, where mining started some 20 years ago. The approach to tailings disposal paralleled the practice for other hydrometallurgical plants treating gold and base metal ores. Impoundment areas were designed to retain solids, and a clear and neutral overflow was considered satisfactory practice. Now experience has shown that these areas, some idle for over a dozen years, release acids, in seepage and overflows, to an unacceptable degree. To protect natural water courses neutralizing plants are operated wherever required. Lime slurry is fed continuously into the tailings outflows in a quantity sufficient to raise the pH to 8.5 and precipitate heavy metals that may be in solution. The objection to this procedure is that the plants will require servicing indefinitely, unless a better remedy is found.

Jan 1, 1975

Power plants that utilize conventional wet flue-gas desulfurization methods produce enormous amounts of flue-gas scrubber sludge. A typical 400 MW power station that burns a high sulfur coal produces 170,000 tons of wet flue-gas scrubber sludge per year. This scrubber sludge is currently landfilled at an enormous cost to utility companies. This material could be used as a raw material for industries that use gypsum, such as the wallboard and plaster industries. However, before it can be used as feedstock for these industries, some of the impurities such as unreacted limestone and silicates must be removed from the scrubber sludge. Froth flotation has tremendous potential for separating these impurities from scrubber sludge. In order to select an optimum flotation process, it is necessary to characterize the surface properties of the major components in the scrubber sludge, which consist of calcium sulfite, calcium sulfate, calcium carbonate (limestone), and silica. In this study zeta potential measurements were used to determine the surface charge of the components in scrubber sludge, with the goal of optimizing a froth flotation process to produce a marketable gypsum product for the wallboard and plaster industries.

Jan 1, 1996

By John L. Burnett

Field criteria are needed to help the prospector recognize a potentially expansible shale, clay, or slate. Certain criteria have proven useful in California even though the physical-chemical mechanisms of expansion are not thoroughly understood. Three conditions must be satisfied before expansion can take place, The raw material must partially fuse at a low temperature (1800-2000°F.), Concurrently a gas must form. The raw material must be sufficiently impermeable to trap the gas within the particle. Low temperature melting, which is controlled by bulk chemical composition, is difficult to assess in the field but most common shale, clay, and slate is suitable in this respect, The presence of organic carbon, the most common source of an expanding gas, is indicated by dark color of the rock, A large proportion of clay-sized particles promotes impermeability; fine grain-size is indicated by conchoidal fracture, The shape of the expanded particle is commonly indicated by the texture of the raw material. Rocks with well-developed slatey cleavage produce accordion-shaped particles; massive rocks tend to form rounded particles. Undesirable materials commonly present with expansible shale include: weathered shale; sandstone and other coarse-grained sediments; and limestone, Weathering, which is indicated by color, reduces or eliminates expansion; therefore, surface samples seldom are representative of the shale found at depth. Sandstone does not expand and will increase the average weight of the finished product, Limestone or particles of carbonate minerals can be actively harmful because they are converted to free lime (CaO) during the firing process, If free lime is present in concrete, hydration will cause expansion of the lime particles which result in surficial ?popouts.?

Jan 1, 1963

By T. Orgodol, P. Korpi

This paper describes the Boroo Gold Company (BGC) ore processing plant which is the first CIL/CIP plant in Mongolia. BGC is the first foreign investment in Mongolia since 1990 and follows the highest safety and environmental standards. Commissioning of the plant commenced in October 2003 and commercial production began in March 2004. Gold recovery is by gravity and conventional leaching and carbon adsorption. Detoxification of the leach tails is by the SO2/Air method before discharging the tails to the tailings pond. Ore types treated at Boroo are predominantly oxide with some transition ore as mining reaches the lower sulphide zone. There are several interesting technological issues and projects ongoing at Boroo which will also be discussed.

Jan 1, 2007

By C. G. Anderson

The treatment of concentrates for the recovery of the precious metals imposes many processing and environmental challenges. Hydrometallurgical technology for recovery of values from refractory precious metal materials has been developed at Sunshine Mining and Refining Company. Historically, conventional pyrometallurgical methods have been used for to recover values from refractory precious metal materials. However, enhanced environmental awareness has worked to constrain typical refractory material processing methods. This hinders and limits the ability of established smelting or roasting technologies to fill the needs of new sources of refractory materials. An alternative is to treat refractory precious metal bearing materials by hydro metallurgical techniques. These methods are considered more environmentally appropriate since they do not produce gaseous pollutants like sulfur dioxide. Industrially, Sunshine Mining and Refining Company employs hydrometallurgical technology to process refractory precious metal materials. The Sunshine antimony leach uses a mixture of sulfur and caustic under ambient pressure and boiling conditions. The Sunshine oxidation leach uses a mixture of sulfuric acid and a minimal amount of sodium nitrite under relatively mild oxygen pressure and temperature conditions. These conditions result in a nitrogen species catalyzed sulfuric acid pressure leach. Since hydrometallurgical technologies provide industrially proven processing options for many types of refractory precious metal containing materials, their usage should be considered. Ample work has been conducted to prove the effectiveness of Sunshine refractory processing technologies on various materials. This has included platinum group metal materials, refractory gold concentrates, silver materials and indium materials. In some cases, complete process flowsheets have been designed to recover both base and precious metals from these materials and to treat the generated waste streams in environmentally sound manners.

Jan 1, 1996

Extensive testing in the past ten years has shown that where a uniform natural gamma background is present in the strata bordering the seam, the thickness of the boundary coal left in place after mining can be determined by measuring the attenuation of that radiation (Nelson and Bessinger, 1989). Measurements made by the authors in underground mines in Pennsylvania, West Virginia, Ohio, Illinois, and Kentucky, and by others in Wyoming and New Mexico, have shown the presence of such a gamma background (Nelson, 1989). Natural gamma coal thickness sensors of several configurations have been tested in mines owned and operated by the Consolidation Coal Company in Pennsylvania and west Virginia (Nelson and Bessinger, 1988).

Jan 1, 1990

By Neal I. Rosenberg

There is an old education maxim that a good teacher "first tells 'em what he's gonna tell 'em, then he tells 'em, and finally he tells 'em what he told 'em. I'm sure this says something about me as a teacher, and therefore as a student as well, but I'm going to begin by violating that maxim. The first thing I'm going to do is tell you what I'm not going to tell you. I'm not going to tell you that my company has found the solution to frozen coal problems. I'm not going to tell you that there is a miraculous new formula, and your frozen coal worries are over. Actually, it is really questionable whether there is a single step or process which will provide the answer. On the other hand, I do believe that at Union Electric we have made significant improvements in our ability to deal with frozen coal over the last couple of years, and I propose to describe some of the things we have done, what conclusions we have drawn so far, and the direction in which we are now headed. It would be well here to give a brief picture of the Union Electric system and the frozen coal problems we face. Union Electric is an investor owned electric utility with headquarters in St. Louis, Mo. Total generating capability is on the order of 6800 megawatts, which snakes us 24th or 25th largest in the United States. Power generation is by steam (both coal and oil fired), hydro, pumped hydro, and combustion turbine units. Although that is a diverse power supply, by far the largest source is our four coal fired steam plants. One receives coal exclusively by barge, the other three by rail. It is at these last three plants that the frozen coal difficulties arise.

Jan 1, 1980

By C. D. Atkinson

The past few years have seen a much greater concern about air pollution in this country. This concern has led to increased legislation restricting the discharge of many substances such as sulfur dioxide and particulate matter to the atmosphere. In the future, it is anticipated the present permissible emission levels will be lowered. Current and impending restrictions on the emission of sulfur dioxide have brought out the fact that there is a shortage of available low-sulfur fuel for power plants and industrial use which promises to have an economic impact on the nation's energy picture (1) The Pittsburg & Midway Coal Mining Co., under the sponsorship of the Office of Coal Research, has developed a process to produce a fuel capable of meeting the energy requirements of this nation (2) The process is called "The Pittsburg & Midway Solvent Refined Coal Process". The principle product from the process is Solvent Refined Coal which is an ashless, low-sulfur fuel with 16,000 BTU per pound regardless of the rank coal from which it is produced. The product can be burned as either a solid or a liquid fuel. The process appears to be economically more attractive than other methods of alleviating the pollution problems of this nation. In addition to reducing the air pollution caused by stack gases, the product produced is potentially useful as a source of electrode coke and electrode binder pitch, as a gas turbine fuel, in the production of metallurigical coke and pavement asphalt, and as the first step in the overall process to produce distillate fuels and pipeline gas (3).

Jan 1, 1973

By Ralph S. Mattson, Frederick T. Hancock

GENERAL DESCRIPTION The Granduc mine is situated on the west side of Granduc Mountain in the Leduc River area of north¬west British Columbia about 51 km (32 miles) north¬east of the town of Stewart. Entry is by means of a 16.6-km (10.3-mile) long haulage adit from the concentrator site at Tide Lake. The principal ore mineral is chalcopyrite. Ore mineralization varies from dis¬seminated to irregular semimassive stringers generally associated with highly folded and sheared biotite-rich sections of metasediments. The ore bodies vary in width from 3.6 m (12 ft) to over 18 m (60 ft) in a zone some 800 m (2600 ft) in length dipping at an average of 75° to the west and plunging steeply to the south. Capping varies in thickness from 0 to over 300 m (1000 ft). SELECTION OF MINING METHOD In the early stages of planning the primary require¬ments to be met by the mining method were: (1) suit¬ability to high output and productivity, (2) adaptability to mechanization in order to minimize operating costs, and (3) minimization of the costs and the development period for preproduction. Considering these, the selection of a mining method was narrowed to a choice be¬tween sublevel long-hole stoping or sublevel caving. A method utilizing backfill was ruled out as uneconomical, especially since mill tailings would not be readily available. At the time, the following facts about the ore body were recognized: (1) the average width of the ore body was 12 m (40 ft), (2) 30% of the stoping blocks were in areas of 6 m (20 ft) or less in width; (3) in many cases, ore zones were located in parallel lenses or were irregular in shape; (4) major faults were present in the hanging wall close to the ore, with minor faults within the different ore zones; and (5) the ore itself was lami¬nated with fractures and jointing crossing such bedding planes. Considering these factors, the two stoping meth¬ods were compared. Sublevel Open Stoping The characteristics of sublevel open stoping are: 1) It is generally not used for widths below 6 m (20 ft) since the ratio of waste development increases as level intervals are reduced to insure good control of long-hole drilling. 2) This method is used where ground is relatively competent. Otherwise more pillar support is required, increasing ore losses or mining costs with pillar recovery. 3) The method is not easily adaptable to parallel ore zones separated by relatively narrow waste bands. 4) Close control of pillar and stope design will per¬mit some degree of dilution control. Sublevel Caving Sublevel caving is characterized by the following: 1) The method can be used for a variety of ore widths, recognizing that 3 m (10 ft) in width may be an economical minimum. 2) It is particularly favorable where wall rock is weak. 3) The method has flexibility so that irregular ore bodies or parallel lenses can be mined. 4) The need to leave pillars for support is eliminated 5) Dilution and tonnage recovery are interrelated. With concern about the stability of the hanging wall and the geometry of the ore bodies, there was a growing favor to choose the sublevel caving method. However, this method was relatively new to Canadian mining practice. Teams were sent to Sweden to tour mines using sublevel caving and to study their methods and any extraction and dilution problems. The objective was to develop cost and performance data and ore recovery grades and tonnages for feasibility comparisons. Next, detailed mine plans were developed covering a three-year period using sublevel caving and sublevel open stoping. Extensive analyses of the development and production costs for each method were made. With sublevel caving, ore losses of 10% with a 20% dilution rate were used. The same ore loss was estimated for sublevel open stoping with dilution varying from 11 to 15%. As a result of these studies, the decision was made to use sublevel caving partly because slightly lower op¬erating costs, together with higher productivity, offset the adverse effects of higher dilution; and partly be¬cause, at the start, there would be less risk in being able to handle the ground, and the method offered more flexi¬bility. Open stoping was not ruled out for testing at a later date. It was recognized that some other method of mining would be required for mining the ore extensions below the glacier, where surface subsidence must be avoided. During the operating life of the mine, open stoping was tested in narrow width ore zones and found un¬satisfactory due to wall failure. A cut-and-fill stope, utilizing run-of-mine waste from development, was also operated on a trial basis. A sequence was finally estab¬lished that permitted ground control; however, produc¬tivity needed considerable improvement in order for open stoping to be considered as a sole alternative for mining below the main haulage level. At the termination of operations in June 1978 over 13 000 000 t (14,500,000 st) had been extracted from above the main haulage level at production rates varying from 3270 to 7250 t/d (4000 to 8000 stpd). Sublevel Caving Methods Two basic methods were used, transverse and longi¬tudinal. The transverse method was used mainly in the upper C ore body in widths greater than 18 m (60 ft), and occasionally as an alternative to the longitudinal method in localized areas where strike faulting inter¬fered with normal development. The longitudinal method was used in ore bodies varying in width from 3.6 to 18 m (12 to 60 ft), with a multilongitudinal varia

Jan 1, 1982

By T. Janke

When Homestake Mining Company's Mclaughlin Gold Project was being evaluated for development, it was apparent that recovery of gold from mined ore would be difficult and relatively expensive. The ore was refractory, averaging less than 50% recovery in conventional direct cyanidation milling tests. Extensive testing narrowed the processing choices to either autoclave or roaster oxidation before cyanidation. Roasting was ruled out due to higher capital and operating costs --the Mclaughlin mill would be the first large-scale application of autoclave oxidation of a refractory gold ore.

Jan 1, 1989

By J. M. Manka

American mining companies operate many drum hoists that hoist from various depths with varying payloads. This paper will review the procedure of selecting optimum wire rope constructions for these existing hoisting installations. Selection for proposed installations along with recommended hoisting parameters will also be presented. To achieve this optimum rope performance, preventative and corrective maintenance procedures will be reviewed along with examples of rope problems and their causes.

Jan 1, 1992

By R. D. Auten

This paper is primarily a report on the process of using oolitic aragonite, as received from Marconi& Ocean Industries, Ltd. without further processing, for a major portion of the raw feed to a rotary kiln producing cement clinker. Oolitic aragonite is a high purity calcium carbonate material, dust free, with the appearance of a course, white sand. It is farmed when under the right conditions of temperature ocean currents, and ocean floor topography. The calcium carbonate will precipitate into small spherical particles on the ocean floor. These oceanic conditions are correct to cause precipitation of the calcium carbonate near Ocean Cay in the Bahama Island chain. The present growing deposit near Ocean Cay is in the billions of tons and is mined by a relatively shallow-operation hydraulic dredge. The aragonite thus formed is en orthorhombic crystaline form of calcium carbonate with a specific gravity of approximately 2.9. At a temperature of 505°C. (940F), the aragonite converts to the hexagonal crystal of calcite with a specific gravity of 2.6 to 2.7. This change in specific gravity weakens the structure of the particles but the individual pieces are still fairly hard. When the material is then calcined at 900°C (1650°F), the particles retain their shape but become soft and will break up under finger pressure.

Jan 1, 1978

By Paul H. Johnson

Heap leaching is a simple means of recovering copper from oxide and oxide-sulfide ores. Although this method requires relatively low capital investment, it suffers from low copper recoveries and relatively high acid consumption. The objectives of this study were (1) to determine what factors influence these shortcomings, and (2) to establish a basis of correlation. between laboratory and field testing. The principal methods of leaching copper ores at the present time are heap, dump, in-place, and vat. A number of references are presented in the bibliography on the theoretical and practical aspects of these methods. In the heap-leaching method, crushed or uncrushed oxidized copper ore is placed on a prepared drainage pad. Acid leach solutions are dis¬tributed over the top of the heap, caused to percolate downward through the ore bed by gravity, and are drained from its base. The leach duration is generally a matter of a few months. Dump leaching is used to recover copper from very low-grade oxide-sulfide pit-waste materials. Here again, leach solutions are, percolated downward through a heap. However, in this method, little or no acid is added to the leach liquors, prepared drainage pads are not used and the leaching cycle is generally measured in terms of several years. Leaching agents are usually autogenously generated by the breakdown of sulfide minerals within the dump.

Jan 1, 1970

By R. M. Geisenheyner

We are in the midst of a transportation revolution. From the Wright Brothers to the SST. From the Toonerville Trolly to the 120-mile-an-hour Northeast Corridor train. From Christopher Columbus to Neil Armstrong. There is another side to this revolution and it is going on every day at our mines and factories. We call it bulk distribution. Perhaps ordinary sounding, but every bit as exciting as those other transportation innovations. Basically, we have gone from bag to bulk, and in the process reached that Last Frontier in manufacturing cost savings. By revitalization of our distribution system, we have made it possible for our mines and factories to realize fully the vast economies of spage-age manufacturing techniques and accomplishments. The pacesetters in this distribution revolution have been you gentlemen, the producers and shippers of dry mineral products. In this area bulk distribution grew up. And here most of the current innovations are being introduced. Management of your industry was early to realize the economic advantages of bulk distribution. Distribution costs of your products comprises a sizeable, portion of the selling price and management is sensitive of any developments which may even slightly reduce the cost of transportation, marketing or the exploration of new frontiers.

Jan 1, 1969

By Joseph H. Herz

Recently, the President of the Portland Cement Association, Robert D. Maclean, gave the keynote address at the Rock Products Eighth International Cement Industry Seminar. He said, after listing the seven plagues our industry is suffering, "none of these items I have listed is a sales or marketing problem. Further, none is a problem that can be relegated to long-term study by public or private research agencies. They are here-and-now problems to be faced daily by producers, researchers, equipment manufacturers and consultants. THE DAY OF THE PRODUCTION MAN IS HERE." (1)

Jan 1, 1973

By D. L. Slyter

Coeur d'Alene Mines Corporation's Thunder Mountain project in Valley County, Idaho, is one of the most remote active mining operations in the US. High altitude, heavy snows, and proximity to wilderness complicate operations in this environmentally sensitive area. Mining is conducted by conventional open pit techniques while processing utilizes on-off heap leaching and Merrill-Crowe zinc precipitation. The project is one of a few similar operations which detoxify the leached ore prior to its removal from the leach pads. Both leaching and neutralizing ( detoxifying) circuits are closed and independent loops. After leaching, alkaline chlorine solutions destroy residual cyanide in the heaps. Calcium hypochlorite solution for the cyanide destruction is prepared by injecting chlorine gas into milk of lime solution.

Jan 1, 1998

By C. K. Lee

The influence of passageway fires on ventilation flows was investigated in three horizontal fire tunnel s: 1. a laboratory-scale fire tunnel network, 2. a small-scale fire-plume tunnel, and 3. a large- scale simulated mine gallery. Wood was used in the first and third tunnel and gas-burners were used in the second one to generate passageway fires. Experimental measurements included gas mass flow rate, velocity, temperature and pressure in the passageways and exhaust fans. The overall and local effects of the fires were investigated by comparing these measurements obtained before and during the fires. Results show that a passageway fire essentially increases the flow resistances of the passageways of a tunnel flow network by virtue of volatile fuel mass injection and high combustion temperatures. The ventilation air flow rate is consequently throttled to lower values than that before the fire. At severe throttling, reverse flow occurs. Conditions for reverse flow are characterized by two Froude numbers based on gas velocity and density upstream of the fire, and height of the passageway. The Froude numbers and the characteristics of the passageway fire are necessary to analyze mine ventilation/fire interaction and to predict mine ventilation patterns during fires.

Jan 1, 1980

By William Hustrulid

INTRODUCTION The Idarado mine lies high in the San Juan Moun¬tains on the divide between the Uncompahgre and San Miguel Rivers and consists of a consolidation of a number of old and prominent mining properties through which course some famous and very produc¬tive veins. Among the better known are the Smuggler, Tomboy, Montana-Argentine, Black Bear, Liberty Bell, Virginius, Flat, Barstow, and Japan, as well as many others. Most of these veins have been extensively mined over the past century and in the last three decades the Montana-Argentine and Black Bear veins have been the backbone of the Idarado mine; therefore, the descrip¬tion of the veins is limited to them. Access to the mine is through either the Treasury tunnel, whose portal is below Red Mountain Pass on US Highway 550 at an elevation of 3244 m (10,643 ft), or the Mill Level tunnel entrance 3.2 km (2 miles) east of Telluride, CO, at an elevation of 2761 m (9060 ft). The Treasury tunnel intersects the Black Bear vein 2643 m (8670 ft) from the portal and the Mill Level tunnel intersects the Argentine vein 2179 m (7150 ft) from the portal. Mining is by shrinkage stoping from slusher sublevels. The size of the scope blocks varies somewhat, but the standard size is 67 to 76 m (220 to 250 ft) long and 61 to 76 m (200 to 250 ft) high. The mine ranks either first or second in Colorado in yearly production of gold, silver, copper, lead, and zinc. It is 9.6 km (6 miles) from the Red Mountain plant to the Pandora plant, via interconnecting drifts and raises. There are engineering offices at both plants as a matter of convenience. The Red Mountain plant in¬cludes the company general offices, warehouse, carpen¬ter and machine shops, and mine change room. The Pandora plant consists of the mill and assay office, ma¬chine shops, and mine change room. The flotation mill has a capacity of 1632 t/d (1800 stpd), making a bul¬lion product and separate concentrates of lead, copper, and zinc. HISTORY OF THE MINE The Montana-Argentine vein was first extensively worked by the Tomboy Gold Mines Co., Ltd., a British concern. This company mined the stoped areas above the Ophir level between 1910 and the late 1920s and most of the stoped areas above the 2100 level between 1900 and the late 1920s. Gold was the principal ore metal mined. The area between the Revenue and Ophir levels was mined chiefly by the Revenue Mines Co. between 1900 and 1910. The ore was worked from the Revenue tun¬nel, which portals in Canyon Creek. Gold and silver were the chief metals recovered. The stopes between the 1700 and Revenue levels, as well as some higher stopes, were mined by Telluride Mines Inc. during the 1940s. The Mill Level tunnel was driven by that company from 1945 to 1948. Lead and zinc then became economically more important than the precious metals. In 1953, Idarado purchased Telluride Mines, which merged with the parent company in 1956. The Black Bear vein was first extensively worked by the Black Bear Mining Co. in the 1900s and by the Colorado Superior Mining Co. from about 1914 until snowslides at the mine camp [altitude 3750 m (12,300 ft)] terminated the company's operations in 1924. Leasers operated at intervals until 1934. The Treasury tunnel, formerly the Hammond tunnel, had been started before 1900 and reached the 1646-m (5400-ft) mark early in the 1900s, at which time activity lagged until the late 1930s. In the early 1940s, Idarado extended the Treasury tunnel from its heading at 1646 m (5400 ft) to the Black Bear vein and established a raise connec¬tion with the 600 level, the lowest level in the old mine. Since completion of initial work in the mid-1940s, systematic development of the mine, both in the driv¬ing of new headings and the utilization of older open¬ings, has resulted in the present extensive network of workings. GEOLOGY Introduction The oldest rocks of the area are the Precambrian metamorphic rocks, the Uncompahgre formation which are massive quartzites, some phyllites and slates, and dolomitic or limestone beds. West-dipping Paleozoic and Mesozoic strata lie on the Uncompahgre formation. These units are separated by a major angular uncomformity from overlying, essentially horizontal, Tertiary formations that include the basal Telluride conglomerate and several thousand feet of overlying volcanic rock. Intrusive rocks, mostly Tertiary in age, are common in the area, and are in the form of dikes, stocks, and sills. Dikes are very closely associated with some of the ore veins (Mayer). Description of Veins The Black Bear and Argentine veins range from 0.6 to 7.6 m (2 to 25 ft) in width, but in most places are 1.5 to 2 m (5 to 7 ft) wide. They vary in character from a well-defined tabular structure between sharp "frozen" walls, or gouge seams, to an irregular zone of quartz and quartz-sulfide stringers. Many gouge seams within the veins make the veins blocky and loose. Common gangue minerals in the veins include quartz, pyrite, rhodonite, chlorite, sericite, clay minerals, epidote, calcite, adularia, rhodocrosite, fluorite, and specularite. Quartz constitutes 60 to 70% of the veins and varies widely in character, ranging in color from clear through white, gray, and green, to amethyst. Chlorite, sericite, the clay minerals, and fine-grained quartz are common alteration products of the vein walls and of wall rock fragments in the veins. Epidote is an altera¬tion of the dike or, less commonly, of tuff-breccia. Calcite, adularia, fluorite, and rhodocrosite, which are wide¬

Jan 1, 1982

By Alexander N. Winchell

The cities of St. Paul and Minneapolis are located side by side on a double loop in the Mississippi River near the eastern boundary of Minnesota. Both cities lie on both sides 'of the Mississippi River, which in this region describes an S-shaped curve. The two cities together have a population of over 600,000 people, gathered in an area which has grown to be a single urban district inasmuch as the two cities have expanded until they occupy all the territory formerly between them, though the business centers are ten miles apart. St. Paul is located at the head of navigation on the Mississippi River and Minneapolis, farther upstream, at the Falls of St. Anthony. A recent water¬power development below the Falls of St. Anthony, constructed by the Federal Government, has at the same time made possible navigation to the foot of St. Anthony Falls through a lock around the dam which replaces the former rapids. Minneapolis, in 1920, had a population of 380,000 people, and St. Paul, 235,000. The two cities occupy land at an elevation of 700 to 900 feet above sea level. The average elevation is slightly less at St. Paul. Minneapolis has nearly 4,000 acres of parks and park¬ways, while St. Paul has nearly 2,000.

Jan 1, 1920