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By Hisaaki Takata, Nobukazu Nanko, Ichiro Izawa

INTRODUCTION Tochibora Mine belonged to Kamioka Mines, which are located in Gifu Prefecture, the cen- tral part of the Japanese main island, has occupied the important position as a main source of silver, lead and zinc for the smelter of Mitsui Mining & Smelting Co., Ltd.. The annual production of crude ore goes up to about 1 mil- lion metric tons with average grades of 20g/t Ag. 0.31% Pb, and 4.18% Zn. For the last 20 years, the mine has struggled with the depressed zinc metal price in yen and the escalation of operating cost, especially la- bour cost. For instance, prices of commodities and labour cost have been raised 8 times and 14 times respectively during this period. On the other hand, total operating cost has been con- trolled to increase 4.5 times. The productivity of the mine has been increased 4 times by means of improvement of individual works and adoption of perfectly mechanized trackless mining system to reform the mining method. Some improvements are presented as follows; Today, drifting is carried out by the tunnelling crew which consists of 9 operators equipped with 5 Mobile-Jumbos, 2 LHD's and a jeep for carrying and loading ex- plosives. In order to get the high utilization of these machines, each crew manages systematically and continuously adjacent 15 faces under the direction of a forceman. As a result, we have accomplished high advance of drift. Furthermore, raising is carried out by a set of the raise-borer in case of long raises like ore- passes and also by the stage blasting method in case of short ones like slot-cuts. Generally speaking, the sublevel stoping method is considered to be one of the best methods for various respects; high productivity, low cost, easy grade control and safety. This method has been introduced in Tochibora Mine since 1950 and is now widely applied to lodes that could not be mined before by this method be- cause of the shapes of them and the conditions of their country rock. Now we are convinced to be nearly accomplished for our method. This report presents some examples of the standard sublevel stoping method and its variations which are practised at the mine. STANDARD SUBLEVEL STOPING Up to 1979 since 1950 when this method was first applied at the mine, 17 million tons of crude ore have been produced by this method, which amounted to 55% of the total production of the mine. During three decades, the productivity has been increased considerably by means of trackless mining system using large-sized machines and etc.. In other words, mining efficiency has been increased from 20 tons/man.shift to 100 tons/man.shift at large stopes. Formerly undercutting was carried out above the drawing drift by jack-legs and small scrapers and then several cone-shaped glory holes were made by stopers between the undercut level and the draw level. But now, a undercut drift is driven in parallel with the haulage drift at the same level of it. And at the undercut drift, drilling upward is carried out by a large-sized machines like the fan drill in order to extract the broken ore directly at the drawing drift connected with the haulage drift. Simplification of the work as well as utilization of large- sized machines enables to get higher productivity. But if the deposit is comparatively wider and the broken ore around the middle part of the deposit cannot be extracted, several drawing drifts are driven to the deposit and in order to minimize dead ore at the final stage of operation, the

Jan 1, 1981

By Georges Caraghiaur

INTRODUCTION A small, independent mining enterprise (SIME) is an independently owned and operated business concern engaged primarily in mineral development, from the exploration to the production stage. By definition, a SIME is not dominant in its field and has no long-term financial backing. It is not a major producer of minerals, nor is it a petroleum company, a government organization or a subsidiary of such organizations. A SIME may or may not be listed on stock exchanges. The SIME is also called small miner, junior mining company, speculative mining company, and penny-stock mining company. The survival and growth of a SIME is most often controlled by an independently minded entrepreneur, who not only conceives and carries out new projects, but is materially affected by the success or failure of these projects. Usually, a SIME does not have a large or constant cash flow. Such a lack of long-term financial backing forces the SIME to meet the challenge of finding capital for its activities on a step-by-step basis. Its survival strategy is thus based on finding funds at the right time, of the right type and in the right quantities. The purpose of this paper is to show that the type of venture in which a SIME is involved affects the difficulty of raising new capital and the choice of possible sources of funds. The Small, Independent Mining Enterprise (SIME) The SIME is not a figment of the imagination of financial speculators. In Canada, there are more than 1,800 small mineral enterprises which spend on the average less than $1,000,000 per year on activity which is directed at bringing forth new mineral supply. In the United States, 64% of the non-fuel mining operations are conducted by nine or fewer men, and the coal industry is characterized by several thousand smaller independent producers, still predominantly family owned or controlled. In Chile, 80% of the mines can be classified as small mines. In the Philippines, approximately 60% of metal mining operations are classified as small mines. The SIME is actively engaged in mining all over the world, and very much so in the Pacific Rim. The Mineral Supply Process The whole process of bringing forth new mineral supply involves a dynamic sequence of activities, ranging from the initial exploration for a mineral occurrence to the production of a marketable product. A typical sequence is outlined in Figure 1. At the grass-roots exploration stage, a series of geological, geophysical and geochemical tests is aimed at locating a mineral occurrence. Especially in the early phases of such activities, the uncertainty (risk) surrounding a venture is high, since only general geological information exists on which to base a judgement of the economic viability of the venture. Such risk is attractive to few investors. As the information-gathering process continues, uncertainty diminishes. Once exploration efforts result in the discovery of a mineral occurrence, i.e. when indications of mineralization and mineable widths are obtained by drilling, there is still uncertainty concerning the extent of mineralization and the probability of having located an economic mineral deposit. The expected return on investment at such a stage may well be negative but the slight possibility of a very high positive return will make a number of financial sources available to an exploration company. At the development stage, much of the uncertainty

Jan 1, 1985

By C. D. Ramsden

During the war years of 1941 to 1945, maintenance of mine locomotives and other mine equipment took the form of rebuilding rather than of renewing. Purchase of new parts became increasingly difficult and mine operators found it necessary to maintain equipment by salvaging old parts. Mine-locomotive wheels, although not representing a high cost per pound of material, contain a considerable amount of steel and without rebuilding would have to be scrapped long before even a moderate percentage of the steel in the wheel is actually consumed. Many mine operators have been rebuilding locomotive wheels for several years with automatic open are-welding equipment. In 1927, the Pittsburgh Coal Co. started such a program, using a single Westinghouse automatic welding head. Since that time, many refinements have been added, until in 1944 the setup was operating very efficiently. A desire to increase wheel life and to hold down rising costs led the company in 1944 to investigate the possibilities of the under-flux are-welding process for wheel rebuilding, although at that time, little or no work had been done on multiple pass or "build-up" welding with the underflux arc. Work with this process had been confined largely to welding together flat plates in a "down hand" position. Possibilities for savings were intensified by the fact that the company has 300 locomotives in operation, using actively or as spares 1200 sets of wheels. During 1942 and 1943, about 500 sets of wheels were sent each year to the company's maintenance shop at Library, Pa., for rebuilding. Based on having 800 sets in service (many of the company's heavy locomotives have three sets of trucks), the average life of a set of wheels was approximately 19 months. Anything that could be done to increase wheel life promised sizable annual savings. Open Arc Welding With the open arc process, a set of wheels could be rebuilt for about 40 pet of the cost' of new wheels. Underflux arc welding gave promise of greatly increased welding speeds, more homogeneous, dense deposits and the chance for use of a wider range in welding-wire compositions. Against this was ranged the higher cost of materials and the cost of developing the process from the theory to a practical reality. Since 1927, when a single open arc head was placed in operation, fixed welding procedures had been gradually developed, a positioner for handling a set of wheels was built, and a second head was placed in operation in order to weld two wheels simultaneously. The fixture, or positioner, which was built to position and turn the wheels under the two welding heads, included a motor drive, controlled by a rheostat, which rotated the wheels through a double set of speed reducers and a removable jaw coupling attached to one end of the axle of the wheel truck.

Jan 1, 1947

By Irving B. Cadoff, Ernest Levine

The fracture path in a Cu-4 pct Ag alloy wet with mercury was found to he dependent on the heat treatment used prior to tensile testing. Both quenched and slow-cooled alloys were embrittled by the mercury.. For quenched alloys, which were single -phase, the crack path was intergranular. Slowly cooled alloys, which contained precipitate platelets, exhibited transgranular fracture. By comparing stress-strain data for these alloys with similar data on mercury -embyittled solid-solution alloys, it was found that the role of the precipitate and that of the grain boundary in the fracture process were equivalent . THE fracture path which results when metals are strained in a liquid environment appears to be dependent upon the nature of the specific system. Mixed transgranular and intergranular cracking has been observed in the 2024 T4 aluminum alloy when wet with a Hg-Zn amalgam.' Transgranular fracture was induced in an A1-4 pct Cu alloy by masking off the grain boundaries from the liquid Hg-Zn amalgam.' The embrittlement of solid-solution alloys of copper in mercury results in intergranular fracture.' The effect of a precipitated phase on the fracture path in copper alloys is the subject of the present investigation. Cu-4 pct Ag was chosen so that its behavior could be compared with previous work on embrittlement of copper alloys. PROCEDURE High-purity copper (99.999 pct) was used as a base metal, along with 99.999 pct purity silver. Melting was carried out in graphite crucibles with manual agitation to get adequate mixing. The ingots were homogenized by means of vacuum annealing, chemically cleaned, and cold-rolled with intermediate vacuum anneals, to the desired thickness of 0.025 in. The rolled sheets were sheared to the final specimen size of 1 by 0.025 by 0.1 in. The specimens were then electropolished in phosphoric acid (concentrated) and heat-treated in an evacuated capsule at the solution temperature of 770°C. One batch was slow-cooled and the other quenched in water. The thermal etching which occurred during heat treatment was sufficient to permit grain boundary and precipitate delineation. Tensile tests were carried out in air and in mercury on a Table-Model Instron at a pulling speed of 0.2 in. per min. Wetting was accomplished by electro-polishing in phosphoric acid, rinsing in alcohol, and immersing in a pool of mercury. Once wet the coating was stable throughout the mechanical tests. In our studies it was found that wetting the surface or total immersion in mercury produced identical results. For convenience, the wetting technique was used for the experiments described. The crack pattern was observed directly on the surface by removing the mercury after testing. De-wetting was accomplished by flame heating the samples in a vacuum after testing and before observation of the surface. No etchant or polishing was used so as not to obscure the crack pattern on the surface. The mercury removal, however, resulted in a pitted surface but otherwise left the pertinent features visible. Photomicrographs were taken in this condition so as not to disturb the structure associated with the crack pattern. The average precipitate spacing was measured previous to coating the sample with mercury by use of an optical eyepiece micrometer. RESULTS The microstructure of the quenched specimens was single-phase while that of the slow-cooled specimens contained thin platelets of a second

Jan 1, 1964

By C. D. Ramsden

During the war years of 1941 to 1945, maintenance of mine locomotives and other mine equipment took the form of rebuilding rather than of renewing. Purchase of new parts became increasingly difficult and mine operators found it necessary to maintain equipment by salvaging old parts. Mine-locomotive wheels, although not representing a high cost per pound of material, contain a considerable amount of steel and without rebuilding would have to be scrapped long before even a moderate percentage of the steel in the wheel is actually consumed. Many mine operators have been rebuilding locomotive wheels for several years with automatic open are-welding equipment. In 1927, the Pittsburgh Coal Co. started such a program, using a single Westinghouse automatic welding head. Since that time, many refinements have been added, until in 1944 the setup was operating very efficiently. A desire to increase wheel life and to hold down rising costs led the company in 1944 to investigate the possibilities of the under-flux are-welding process for wheel rebuilding, although at that time, little or no work had been done on multiple pass or "build-up" welding with the underflux arc. Work with this process had been confined largely to welding together flat plates in a "down hand" position. Possibilities for savings were intensified by the fact that the company has 300 locomotives in operation, using actively or as spares 1200 sets of wheels. During 1942 and 1943, about 500 sets of wheels were sent each year to the company's maintenance shop at Library, Pa., for rebuilding. Based on having 800 sets in service (many of the company's heavy locomotives have three sets of trucks), the average life of a set of wheels was approximately 19 months. Anything that could be done to increase wheel life promised sizable annual savings. Open Arc Welding With the open arc process, a set of wheels could be rebuilt for about 40 pet of the cost' of new wheels. Underflux arc welding gave promise of greatly increased welding speeds, more homogeneous, dense deposits and the chance for use of a wider range in welding-wire compositions. Against this was ranged the higher cost of materials and the cost of developing the process from the theory to a practical reality. Since 1927, when a single open arc head was placed in operation, fixed welding procedures had been gradually developed, a positioner for handling a set of wheels was built, and a second head was placed in operation in order to weld two wheels simultaneously. The fixture, or positioner, which was built to position and turn the wheels under the two welding heads, included a motor drive, controlled by a rheostat, which rotated the wheels through a double set of speed reducers and a removable jaw coupling attached to one end of the axle of the wheel truck.

Jan 1, 1947

By W. A. Wood, H. L. Wain, R. I. Garrod

The mechanical behavior of recrystallized chromil~m of high purity has been studied, principally in torsion and to a lesser extent in tension, at temperatures between —196oand 350oC. Depending upon the conditions of test, the material may show brittle or ductile behavior or an intermediate or subduc-tile state having some of the characteristics of each. This latter state existed at temperatures between approximately ±50°C and was characterized by a) extensive creep at stresses well below the normal fracture stress, b) nonelastic loading/unloading effects during creep, c) strain softening under cyclic loading, and d) final failure of a brittle type. Suggestions are made on the dislocation processes and interactions which give rise to subductile behavior The ductile-brittle transition in bee metals has been studied mainly by tension and bend tests. These methods, however, favor crack propagation which may follow too closely on initiation for detailed study of the transition itself. For work of this nature, therefore, other modes of deformation less favorable to crack propagation may be preferable. This paper describes a study of the ductile-brittle transition in chromium which is based mainly on torsional testing although a few subsidiary experiments were made in tension. It appears to reveal at the transition an intermediate state that cannot properly be described as either brittle or ductile. Torsional tests on chromium have been made by other workers1 but not, as here, primarily for study of this intermediate state. EXPERIMENTAL The chromium, in the form of swaged rod, was prepared by techniques already described.' The particular material used was one in which the main impurities were 0.03 wt pet 0 and 0.006 wt pet N and one in which transition from ductile to brittle states in torsion conveniently fell at about room temperature. For torsion, test specimens were from 5/16-in.-diam rod to provide test portions 1-1/4 in. long by 7/32 in. diam, a twist of 1 deg thereby corresponding throughout this paper to 0.0015 surface shear. The specimens before test were annealed for 2 hr at 1200°C in vacuo, which left them with a re-crystallized grain size of about 0.01 mm, and then electropolished. They were tested under dead-weight loading in an apparatus in which they could be surrounded by a furnace or cooling chamber, tests in practice being made at various temperatures between -196" and +350°C. For tension tests, cylindrical specimens were prepared in a similar manner and tested at room temperature under dead-weight loading. RESULTS (TORSION) Subductile State. The state described above as intermediate between brittle and ductile may be perhaps best defined by reference to torque/twist curves at various temperatures. These are grouped in Fig. 1. Below about -50°C the material was brittle, in the sense that it exhibited little plastic deformation under stress and fractured by cleavage along planes of maximum resolved tension, which in torsion produces the well-known helical type of fracture. The deformation is typified by the curves for -196" and -150°C in Fig. 1; it shows a slight but progressive deviation from linearity above a certain torque, here about 50 lb-in., until fracture occurs at about 95 lb-in. A feature to which it is desired specially to call attention is that under none of the loads did the specimens show significant creep.

Jan 1, 1964

By W. Hurst, A. F. van Everdingen

A distillate-bearing sand complex is often made up of sand stringers of different permeabilities. In order to take into account the influence of the different permeabilities, "parallel flow" is defined, and this concept is used to compute the recovery efficiency of recycling such a sand complex. In the first approximation the position of the wells with respect to one another is not considered. In order to compute the effect on the recovery efficiency of the position of the wells, the area of a field is considered to be made up of rectangles; each rectangle containing one input well in the center and two producers located symmetrically with respect to the input well. Formulas are given from which to obtain the pressure and streamlines in a steady-state flow. A method is devised to compute rapidly the successive positions of the dry gas front. If the field is irregular in shape, pressure and streamlines can be obtained by a potentio-metric electrical model study. By using the concept of "parallel flow" and the results obtained by mathematical analysis or by electrical model study for a homogeneous formation, the effects of different permeabilities and of the location of the wells can be established. From these studies it is apparent that the efficiency of recycling can be profoundly influenced by differences in permeability among the various parallel strata. Introduction With the discovery of an increasing number of gas-distillate fields in the Gulf Coast, engineers have had occasion to give more thought to the performance of these reservoirs, as evidenced by the literature published in recent years. In the majority of studies published so far the formations are considered to consist of uniform, homogeneous sand. In this paper the authors wish to bring out the effect of variations in the permeability of the formations on the over-all recovery efficiency of wet gas in cycling operations. The importance of the permeability profile is clearly indicated by the fact that in gas cycling, highly permeable stringers in an otherwise tight sand can allow complete displacement of wet by dry gas in these stringers long before the wet gas in the tight portions of the sand is displaced. This by-passing will be reflected by a reduction in the condensate content of the production sooner than can be anticipated from a mathematical analysis or an electrical model study on uniform sands. Electrical model studies or mathematical analyses cannot take into consideration the manifold variations of permeability within the reservoir. Experimental as well as analytical difficulties make it necessary to consider the prototype for either of these studies as a perfectly uniform sand. Mathematical analyses are applicable only when a reservoir can be divided into units of simple geometrical form, of uniform thickness throughout. Electric model studies are capable of giving information on reservoirs of any shape or thickness, but it is essential that the sand be considered homogeneous, so that three-dimensional flow can take place. The method of approach in the following

Jan 1, 1946

By Michael Chender

INTRODUCTION The project financing described elsewhere in this book is available to major mining companies on the basis of a strong balance sheet, a significant equity position in a project of proven feasibility, and a demonstrated ability to bring ore deposits into production. "Junior" mining companies, which are most active in Canada, Australia, the US, and South Africa, by definition do not meet these criteria. In many cases, junior companies will need to raise money simply to cover the costs of acquiring and exploring potential deposits. In what follows we will look primarily at junior financing practices in the US and Canada, although the same general principles will apply in Australia and South Africa. The principal vehicles for junior mine company financing are limited partnerships, the private placement of securities and debentures, and public offering of securities, and property deals with stronger companies, in the form of leases, earn-ins and co-ventures. The cost (in terms of dollars or equity) and, to some extent, the applicability of using these difficult approaches varies with the stage of development of the property and with the corresponding degree of risk to the investor. Much junior company fund raising is, in fact, "interim" financing in that it allows the company to improve its properties and develop its assets to the stage where more favorable financing arrangements may become available. Companies will therefore often use several of the approaches described above during the development life of a specific property. LIMITED PARTNERSHIPS Limited partnerships and syndications are a popular way of structuring the initial private financing of mine developments because of their attractive tax consequences for the well-to-do investor. Costs of exploration and development can in many countries be passed through as deductions to the limited partners, who are thus able to gamble on a possible windfall at little real cost. In addition to private Canadian funds, much German money flowed into Canadian exploration ventures on this basis, but stopped in 1982 due to economic conditions in West Germany. In the US, recent regulations have, in the great majority of cases, taken away the ability to "shelter" income in mining limited partnerships. Deductions taken by the individual investor cannot exceed the amount of money he actually has at risk. Nevertheless, the allowable deductions are still attractive and some of the major mining companies are now beginning to look at limited partnerships as a vehicle for attracting additional exploration funding. Ranchers Exploration and Development has successfully put together several drilling funds on this basis. THE PRIVATE PLACEMENT Initial funding for the acquisition and preliminary exploration of prospects is often accomplished through a private placement of unregistered securities or, where it is not a problem to carry debt on the statement, by a private placement sale of debentures. A fledgling company wishing to do a public offering may well be advised by an underwriter to accomplish a small private placement as a precondition of the larger offering. In this way, the sale of the registered securities is not inhibited by the company starting out with a debt-ridden balance sheet and no demonstrated ability to secure the additional financing generally necessary beyond the proceeds of that offering. In the US, the securities offered in a private placement are not registered with the Securities and Exchange Commission (S.E.C.) but are restricted as to their resale, and could be further restricted by a underwriting agreement

Jan 1, 1985

By W. C. Lilliendahl, G. Meister

Uranium metal with a purity of about 99.9 pct was produced on a large scale by fused salt electrolysis with a material efficiency of about 90 pct. The material efficiency depends mainly on bath composition, electrolysis temperature, electrode smothering, and elutriation during washing of the powder. RESEARCH and development work on fused salt electrolysis carried on by the Westinghouse Lamp Div. Bloomfield, N. J., was done under the Manhattan District Corps of Engineers and later under the Atomic Energy Commission for the commercialization of uranium production for the first atomic pile at the University of Chicago in 1942. From the conception of the atomic bomb project it was recognized that metals with impurity levels having little metallurgical significance would be required. Permissible impurity levels were based on neutron absorption. Of particular significance was boron, which could not be present in amounts exceeding 1 ppm. The electrolytic method as developed in 1930 by Driggs and Lilliendahl,1,2 and later improved, yielded uranium of very low boron and other deleterious impurity content. No other method was capable at that time of producing uranium metal of the required purity.' The low boron content of uranium produced by the electrolytic process resulted from the availability of the reactants which were low in boron content, and from the inherent elimination of boron in the fused salt bath. Since the heart of any electrolytic process is the electrolysis cell, this paper will deal with the development of the original bath composition for the efficient production of uranium from a batch process to one of continuous operation. Emphasis also will be placed on other factors which will affect the material efficiency of the process. With the laboratory cell the material efficiency of uranium powder recovery in the early electrolysis was relatively low, about 35 pct based on the ratio of uranium recovered to uranium added. This early process also had other defects. The electro- lytic bath could not be used continuously. After three runs from a freshly charged crucible, the deposit showed a tendency to slip off the electrode as it was raised from the bath. This condition necessitated pouring the charge every fourth run and preparing a new bath. It was assumed at that time that this condition was caused by the increasing fluoride concentration with each addition of uranium salt, which adected the mobility of the uranium ions. Experimental The electrolysis cells differed in size only. They were of the floating cathode type. The laboratory cells were 21/2 in. ID and 6 in. deep and operated at 30 amp and 5 volts. Production was about 150 g metal powder per 8 hr. The plant cells (5 units) were 9 in. ID x 21 in. deep and operated at 1800 amp at 12 volts. The total production unit was designed to produce 250 lb of uranium powder per 8 hr day. Because of the tendency of uranium fines toward spontaneous combustion, material efficiencies were based only on the recovery of useful coarser metal which could be safely handled. The reported efficiencies are of particular significance only as related to the same group of experimental procedures. At the start of the project no changes were contemplated until previous data could be definitely confirmed. Therefore, the first experiments were with the 50 pct CaCl2-50 pct NaCl mixture. The material efficiency of uranium deposition as a function of the number of electrolyses is shown in Table I. In Table I, the decrease in efficiency at 800°C with number of runs is apparent and confirmed the earlier data. Operating at 900°C enabled the number of runs to be doubled, but before pouring the charge the efficiency of recovery became poor. This test indicated that the improved efficiency might have been caused by an increase in the fluidity of the bath.

Jan 1, 1958

By L. L. Newman

As soon as the Congress recognized the alarming rate at which our domestic oil resources were being depleted during the war, it took action to step up the rate of research and development which the Bureau of Mines had been conducting for several years on the production of synthetic liquid fuels from coals. Under Public Law 290, approved April 5, 1944, funds were authorized for the Bureau to proceed with research, development, design and construction of demonstration plants for the hydrogenation of coal, the synthesis of oil from carbon monoxide and hydrogen, and the distillation of oil from shale. During the latter part of 1944, when it became evident that the collapse of Germany was near, a Technical Oil Mission was organized to study the Ger- man techniques of synthetic oil production. Since the cost of hydrogen or synthesis gas is one of the major factors of expense in the operation of hydrogenation or synthesis gas processes, the Bureau of Mines considered it of prime importance to include on the mission gas engineers whose instructions were to examine closely the methods employed by the Germans in solving the problem of providing low- cost hydrogen or synthesis gas for their very extensive operations on synthetic liquid fuels. The Solid Fuels Mission and the Utilities Mission, as well as the Army and Navy Missions, also included teams for the study of German gas production. Although their objectives naturally differed from those of the Technical Oil Mission and from each other, it was recognized that there was considerable overlapping of their fields of investigation. Duplication of effort, however, was largely avoided through the close cooperation between the group leaders of these missions. For example, although I went to Europe as a member of the Technical Oil Mission, my first German assignment was actually made by Dr. Harry F. Yancey, leader of the Solid Fuels Mission- to investigate the brown-coal gasification methods in central Germany. The art of gasification in the United States and England, and to a large extent in the proximity of the coking-coal regions of Germany and the rest of continental Europe, depends essentially on two stages of operation; namely, carbonization and intermittent water-gas generation. As far as the United States is concerned, complete gasification in a single stage of operation is rarely practiced, except in gas-producer operations where the product is a relatively lean gas consumed close to the point of manufacture, and for water-gas production with anthracite as a generator fuel. Complete gasification of bituminous coal has been the subject of numerous investigations. Many relatively simple modifications of the water-gas generator have been tried to permit the direct gasification of low-cost bituminous coal instead of high- cost coke and anthracite for the production of water gas. Although varying degrees of success have been achieved and the water gas produced generally has been higher in calorific value, in only a few isolated in- stances has there been any sustained use of bituminous coal for water-gas production for public utility uses after the shortages of anthracite and coke have been lessened.

Jan 1, 1946

ONE of the largest hard rock operations ever undertaken on the continent is part of the Aluminum Co. of Canada's gigantic British Columbia development in the mountainous wilderness back of the north coastal port of Prince Rupert. To provide electric power for the huge aluminum smelter under construction at Kitimat, 400 miles northwest of Vancouver, Alcan is reversing the flow of a 5000 sq mile drainage area by blocking the eastward-flowing Nechako River with the largest sloping clay core dam in the world. The backed-up waters of a 110-mile chain of lakes will be tapped at their eastern end by a 10-mile tunnel piercing a mile-high mountain range. From this tunnel water will pass through penstocks to a subterranean powerhouse after a drop of 2600 ft and empty into the Kemano River almost at tide water. A projected battery of eight turbine-driven generators will produce more than a million horsepower, although, initially, three generators of 140,000 hp each will be installed. Power will travel 50 miles across another mountain range, over a glacier, and across an arm of the ocean to what will ultimately be the world's largest aluminum smelter-a plant 2 ½ miles long. Ore smelted at Kitimat will come from Jamaica, but carving the tunnel, penstocks, and powerhouses cavern out of the heart of a mountain is so vast a job that most of the problems of a major mining operation are encountered. The powerhouse cavern alone is a big hard rock job. This chamber lies 1400 ft back from the mountain face and will eventually be 1100 ft long, 80 ft wide, and an average 118 ft high. Now under construction, the roof of this cavern will take the form of a parabolic arch. The prime contractors, Morrison-Knudsen Co. of Canada, are to excavate a length of 700 ft by May 1954. This section will provide a main control room, erection bay and space for eight generating units. The roof arch is to be lined with reinforced concrete and a plan now under study provides for concrete curtain walls separated from the rock sides of the cavern to prevent moisture accumulation. Cavern Excavation Excavation of the cavern began with the driving of an 8x8-ft exploratory drift 1500 ft long. Two horseshoe-shaped access and tailrace tunnels were bored to serve excavation work. These tunnels are approximately the same length as the exploratory drift and 25 ft high. The floors of both tunnels were graveled to permit the use of rubber-tired equipment. Eventually, the tailrace tunnel will be enlarged to 40 ft height to drain water from the turbine pits below the powerhouse floor. When the tunnels reached the cavern location, a mucking tunnel, approximately 25 ft square, was driven the projected length of the powerhouse at the floor level of the cavern. Another tunnel, 10 ft square, was carried in the same direction from the exploratory drift at the level of the arch spring line, 65 ft above the cavern floor. To connect these two tunnels, seven shafts 10 ft wide and approximately 80 ft long were driven through to the lower mucking tunnel, 65 ft below. These shafts were located on 120-ft centers, the length of the powerhouse cavern. Blasting is done from this upper tunnel to the final line of the roof arch. Two bulldozers, an International TD-24 and a TD-9, doze blasted material to the shafts where it drops to floor level. From here, the rock is loaded by Eimco loaders and power shovels into rubber-tired Koehring Dumptors and Euclid end-dump trucks. The arched roof section will be completely excavated and concreted before vertical blast holes are drilled down from the level of the arch spring line to the floor of the cavern. The arched roof section with its parabolic curve will have a width of 80 ft and a rise to the crown of 47 ft.

Jan 1, 1952

By J. S. Ll. Leach, L. R. Rubin, M. B. Bever

The energies of formation of ordered and disordered solid solutions of composition CusAu and the energy of ordering in this alloy were determined by tin solution calorimetry. The degree of order was measured by X-ray diffraction and electrical resistance and microhardness measurements were made on ordered and disordered specimens. AMONG the phenomena associated with the order-disorder transformation of a solid solution, the change in internal energy is of special interest because of the part it plays in the various theories of ordering. Published values for the decrease in internal energy accompanying the formation of a superlattice from a disordered solid solution of composition CuAu range from —370 to —2260 cal per gram-atom. Some of these values represent calculations based on theory and others are the results of experimental measurements. The distinction between the change in internal energy, AE, and the change in enthalpy, AH, can here be neglected, because they are approximately equal for solid-state reactions at normal pressure. An analysis of ordering by Bragg and Williams' predicts an energy change of —605 cal per gram-atom for the formation of a superlattice in the alloy Cuau from a completely random solution. Peierls" application to Cuau of Bethe'sb earest-neighbor theory yields —560 cal per gram-atom for the formation of a superlattice from a matrix which initially contains short-range order. Cowley' extended the nearest-neighbor approach to include as many as five shells of neighbors; on this basis a change in energy of —500 cal per gram-atom is expected. Eguchi," using a quantum-mechanical treatment, calculated a value of —2260 cal per gram-atom for the difference in the energy of completely disordered and completely ordered Cu,Au. Sykes and Jones- eated a completely ordered alloy and measured its heat capacity as a function of temperature. This measured heat capacity agrees closely with the corresponding value found by the Kopp-Neumann (or mixture) rule up to about 250°C and above this temperature exceeds it, especially near the critical temperature for ordering. The difference between the integrals with respect to temperature of the observed and the Kopp-Neumann heat capacities was considered to be the energy of ordering. By this method Sykes and Jones found a value of —530 cal per gram-atom. This value is not adjusted for the short-range order remaining above the critical temperature. The pres- ence of such short-range order is suggested by the difference between the measured heat capacity and the extrapolated Kopp-Neumann heat capacity immediately above the critical temperature. Values reported by Weibke and von Quadt' and by Hirabayashi, Nagasaki, and Maniwaa were obtained in the course of investigations primarily aimed at other objectives. Weibke and von Quadt measured the temperature coefficient of the electromotive force of a Cu-CuAu cell. They obtained a value of —1010 cal per gram-atom for the heat of formation of the alloy at 500°C, at which temperature there is no long-range order. They also obtained —1380 cal per gram-atom as the heat of formation of the ordered alloy at 370°C. Considering the heat of formation of the disordered alloy to be independent of temperature, they estimated the energy of ordering at 370°C as —370 cal per gram-atom. At this temperature long-range order is incomplete and the degree of order changes rapidly with temperature. Hirabayashi, Nagasaki, and Maniwa," using an annealing calorimeter, investigated an alloy containing 23.4 rather than 25.0 atomic pct Au and thus could not obtain complete order. Thelr value of the energy of ordering was —490 cal per gram-atom. Orianis has recently investigated the Au-Cu system by the galvanic emf technique. He reports values for the heats of formation of Cu-Au alloys, from which the heat of formation at 427 OC of an alloy of composition CuAu may be found by interpolation. This value is —1080 cal per gram-atom. In the work here reported, disordered and ordered alloys of composition CuAu and corresponding mixtures of gold and copper were dissolved in liquid tin and the heat effects measured. These heat effects are small, since the dissolution of gold in tin is exothermic and the dissolution of copper is endothermic. The method, therefore, yields fairly precise values of the heats of formation of disordered and ordered alloys and of the energy of ordering. Experimental Procedure The calorimeter consisted of a long-necked Dewar flask immersed in a constant temperature salt bath and has been described by Ticknor and Bever." The chief changes in this equipment were an improvement in vacuum and the replacement of the mercury thermoregulator by a resistance thermometer control circuit. The solvent, which was maintained at a constant temperature near 350°C, consisted of 500 grams of 99.99 pct pure tin. The solute samples were mixtures of gold and copper in the proportion corresponding to the composition Cu,Au or solid solutions

Jan 1, 1956

By J. F. Pierce, S. M. Enterline

Since January 1959 a zinc refiner of novel design has been in operation at Blackwell, Okla., producing 99.995+ zinc from the output of the Blackwell horizontal retort smelter. The refiner is a continuously operated, pyro-metallurgical unit with principal dimensions horizonhl permitting one floor operation. It is very ruggedly constructed. Furthermore it is not subjected to deterioration from high concentrations of iron in the feed metal because the boiling units are separate from the rectification columns and constructed with silicon carbide brick arches through which the heat is transferred radiantly to the zinc. A zinc refiner of novel design has been in operation since January 1959 at the Blackwell, Okla. plant of the Blackwell Zinc Co., Inc., a wholly owned subsidiary of American Metal Climax, Inc. Developed by the company's technical and operating staffs, the new refiner produces special high-grade zinc (nominally 99.995+ pct Zn) from the prime western1 slab output of the Blackwell horizontal retort smelter. The refiner is a continuously operated, pyromet-allurgical, high-capacity unit employing fractional distillation to separate zinc from its impurities, consisting principally of lead, cadmium, and iron. In prime western slab, these elements may vary in concentration from 0.005 to 1.6 pct. Table I shows the impurities usually encountered in the Blackwell prime western metal fed to the refiner together with their boiling points and vapor pressures at 907oC. The commercial pyrometallurgical smelting processes for extracting zinc from ore include distillation from retorts and a limited degree of refining can be accomplished by redistillation in similar retorts. To achieve special high-grade purity, an impractical number of batch redistillations would be required, however, with an accompanying loss of much metal. A continuously operated and integrated unit is required in which fractional vaporization and condensation can take place in a large number of stages. The first commercial continuously operated unit was developed by the New Jersey Zinc Co.4 In the separation of any two substances the degree of separation is limited if a mixture exists having a constant boiling point. The Zn-cd5 and Zn-pb6 systems contain no constant boiling points and hence these metals may be separated to any degree of purity. Basically, the pyrometallurgical refining of zinc involves three steps: a) iron, copper, and other metals having no appreciable vapor pressures at 907ºC concentrate in the bath when the impure zinc is boiled, b) lead, bismuth, antimony, and other metals having boiling points above but closer to that of zinc are selectively condensed from the vapor boiled from the bath, and c) cadmium and other metals having lower boiling points than zinc are fractionally distilled by reboiling and subsequent condensation of the zinc from the vapor. The boiling and condensation of zinc must be conducted in vessels constructed of materials which are refractory at the temperature involved (907 ºC) and it is highly desirable that they also be resistant to attack by the various metals. This requirement dictates construction from relatively small, simply shaped pieces which limits the pressures which may be employed to avoid leakage through the numerous joints. Stresses induced must be limited essentially to compressive ones and suitable provision must be made to permit thermal expansion of the components in order to minimize these stresses. Where heat is to be added for boiling or removed for condensation, refractory material through which the heat flows must have a high thermal conductivity. Silicon carbide is highly conductive but is subject to attack by iron and is expensive compared to other commonly employed refractories. A hard, dense aluminous fireclay is well suited to contain these metals where high thermal conductivity is not required. Zinc will penetrate the pores of fireclay brick but only to a limited depth in an exterior wall and has little effect other than an increase in weight and conductivity. The design objectives of the Amax refiner were: a) to achieve a long life by heavy, rugged construction and by employing silicon carbide refractories only where required for heat conductivity and not in contact with liquids of high iron concentrations, b) to simplify operation, manning and control by essentially horizontal construction, and c) to obtain high capacity in a single unit. DESCRIPTION Flow of metal through the Amax refiner is shown diagrammatically in Fig. 1 and schematically in Fig. 2.

Jan 1, 1963

By D. R. Wieland, A. N. Barron, A. R. Hendrickson

A definite relationship has been found between the reactivity of flowing hydrochloric acid and its shear rate in a carbonate fracture. Both flow velocity and fracture width affect the acid reaction rate. Laboratory studies were conducted on acid reactivity at different flow velocities through horizontal-linear fractures, using 15 per cent hydrochloric acid at 80' F and approximately 1,100 psi. Fracture width varied from 0.02 to 0.20 in. These data provide a new basis from which the spending time and penetration of the acid can be estimated. Equations were derived expressing the relationship between injection rate, fracture width, acid concentration, time and fracture height, for linear and radial fracture systems. Because the penetration of the acid before spending is closely related to the extent of productivity increase resulting from an acidizing treatment, these data provide a valuable insight into some of the controlling factors that must be taken into consideration during treatment preplanning. INTRODUCTION Acidizing of carbonate reservoirs to improve production characteristics has been widely practiced since 1932. Originally, it was assumed that the acid uniformly penetrated natural formation pores and flow channels, enlarging them and thereby increasing their flow capacity. Little consideration was given to the reaction rate of the acid, or how far it would penetrate away from the wellbore into the formation, before spending. It has been shown recently' that, unless fractures are present in the rock, very little penetration is attained before spending, and the benefits of the acidizing treatment are largely confined to the immediate vicinity of the well-bore. Therefore, acid treatments may be classified into two categories: (1) matrix acidizing, in which the acid flows through multiple: formation pores; and (2) fracture acidizing, in which the bulk of the acid travels through fractures in the rock, whether natural or induced. When acidizing treatments are conducted at pressures of sufficient magnitude to open and extend such fractures, it is often desirable to inject a propping agent to hold the fracture open after the treating pressure has been released, thus providing a highly conductive flow channel through the rock. In some cases where acid attack produces surface irregularities, the resulting flow passages can be sufficient to provide high conductivity without use of a propping agent. During injection, the acid dissolves the carbonate rock with which it comes in contact until it is spent. Deeper penetration of the spent acid into the formation produces no appreciable benefit (if no propping agent is used) because the unetched fracture faces will rejoin when treating pressures are released and the fracture will "heal", with negligible resultant conductivity. The spending time of the acid thus becomes important in determining how far from the wellbore the improved-conductivity zone extends. The spending time of acid after injection into carbonate rock depends on the rate at which the acid reacts with the rock. This in turn is controlled by a number of factors, as previously reported. 2-4 These include temperature, pressure, acid concentration, rock composition, injection rate and the area-volume relationship between the acid solution volume and the surface area of the formation flow channels through which it penetrates. Thus, in matrix acidizing, where an extremely large area is exposed to the acid, spending is rapid. In contrast, spending time is prolonged in open fractures, where the area-volume ratio is much lower. Thus, greatest penetration of the acid before spending will be achieved in acid-fracturing treatments where fractures are held open by hydraulic pressure. Perkins and Kern5 have presented equations for fracture-width determinations which, for hard formations such as limestone or dolomite, indicate that high injection rates and viscous fluids are required to create wide fractures. Increasing the injection rate will, in itself, produce faster reaction rates. Because this type of reaction can be considered to be first-order diffusion-controlled, '-' the rate of movement of fluid past the rock surface affects the thickness of the diffusion layer, in turn affecting the reaction rate. Staudt, et al,10 report the results of tests in which a rotating marble cylinder was immersed in acid, and the weight loss at different speeds determined. Such test conditions, however, are not truly indicative of the situation in a formation during an acid-fracturing treatment. This paper reports reaction-rate data obtained under

By E. A. Slover

THE usual reverberatory system of smelting cop--1- per concentrate or calcine has for its component parts a furnace and one or two waste heat boilers. These parts are operated on a basis of compromise, since the furnace can send gas to the boilers at too high a temperature and the boilers by plugging, due to dust or slag, can place a definite limit on the amount of fuel the furnace can burn. Over the years the copper concentrate smelting furnace has had few advances in design. The simple rules of design such as the flame should wipe the bath and the speed of the gases should be reasonably low for dust carrying purposes seem to cover the main features. In the construction of the individual furnaces some innovations are always being introduced. Among these are charging so that the work of smelting is a complete bath process, the use of suspended brick arches in place of sprung arches, the use of basic brick, not only in the crucible, but also in roof and sidewalls, the use of various means to feed the charge, the use of magnetite or other heavy material to construct the hearth, water cooling of bridgewall and slag skimming bay, the smelting of raw charge instead of calcine, the use of preheated air, and possibly the use of oxygen-enriched air for combustion. But the general outlines of the furnaces have not changed much except as to size. Furnaces at Hurley As shown on Fig. 1, the furnace at Hurley is 126 ft long between the longitudinal buckstays and 32 ft wide at the skewback plates. The foundation is a concrete retaining wall with piers at intervals that go deeper into the earth. Purposely the wall at the burner end of the furnace is not backed-up as tightly as the other parts of the foundation so that movement due to expansion may take place here rather than into the boiler foundations. Within these foundation retaining walls of concrete, the earth has been removed to allow the placement of the crucible brick base inside of which a silica hearth is laid 4 ft 6 in. in depth. No expansion is left in the brick base and crucible where they are in contact with the hearth. The hearth itself is of quartzite crushed to 1 in. size with fines left in the product. An 8 in. layer is laid and tamped with paving tampers to about 6 in. in thickness. Then a layer of silica flour is spread and vibrated into the hearth. This operation is repeated until a depth of 4 ft 6 in. is occupied by the silica mass onion-skinned in layers of approximately 6 in. Before firing the entire hearth is covered with broken slag to a depth of 4 in. so that a seal may be formed on the hearth. The crucible is completely faced with magnesite chemically bonded brick while the outside, against the foundation, is made of silica brick. The side-walls are carried up with silica brick in which expansion joints are left at intervals. Above the crucible the sidewall is corbelled to form a shelf on which the charge may build up along the side-walls, see Fig. 2. The arch of the furnace is sprung 20 in. silica brick, with the longitudinal centerline horizontal the length of the furnace, and some 9 ft in the center above the bath. Both straight and wedge brick are used in the construction and a thin silica mortar is troweled for joints. After the arch under heat has assumed its permanent shape, a silica slurry is spread over the arch to fill any cracks that have formed, thus giving bearing surface to the brick and preventing dust from entering the body of the arch to act as a future fluxing agent. The uptake of the furnace slopes up to the boiler entrance where a brick pilaster divides the gas stream for the two boilers. Over this flared uptake is a suspended flat arch of firebrick. The pilaster and sidewalls are constructed of firebrick but the bottom of the uptake is lined with silica brick and fettled through holes in the roof with siliceous fettling. Close to the entrance of each boiler is a brick covered slot through which water-cooled dampers may be lowered in event of boiler trouble. These water-cooled dampers are hung permanently in position ready to be lowered when needed. Flexible hoses to follow the dampers as they are lowered are connected at all times and individual chain blocks are used to lower the dampers. A pump supplying water is started before the dampers enter the heat. Charging of the furnace along the sidewalls for some 80 ft from the bridgewall is accomplished by electric vibrating conveyors fed by belt from charge storage bins above the furnace. These conveyors

Jan 1, 1954

By Paul A. Beck

THE occurrence of the (100) [lOO] or "cube" texture upon annealing of cold-rolled copper has been much investigated.' The conditions favorable for its formation were found to be a high final annealing temperaturez or long annealing time," a high reduction of area in cold rolling prior to the final anneal,' and a small penultimate grain size." The effects of penultimate grain size and of rolling reduction were found by Cook and Richards4 to be interrelated in such a way that any combination of them giving lower than a certain value of the final average thickness of the grains in the rolled material leads to a fairly complete cube texture with a given final annealing time and temperature. Also, according to the same authors, at a higher final annealing temperature a larger average rolled grain thickness, i.e., a lower final rolling reduction, is sufficient than at a lower temperature. These somewhat involved conditions can be understood readily on the basis of recent results obtained at this laboratory. Hsun Hu was able to show recently by means of quantitative pole figure determinations that the rolling texture of tough pitch copper, which is almost identical with that of 2s aluminum: may be described roughly as a scatter around four symmetrical "ideal" orientations not very far from (123) [112]. In the case of aluminum, annealing leads to retain-ment of the rolling texture with some decrease of the scatter around the four "ideal" orientations, and to the appearance of a new texture component, namely the cube texture." A microscopic technique, revealing grain orientations by means of oxide film and polarized light, showed that the retainment of the rolling texture is achieved through two different mechanisms operating simultaneously, namely "re-crystallization in situ," and the formation of strain-free grains in orientations different from their local surroundings, but identical with that of another component of the rolling texture. Thus, a local area in the rolled material, having approximately the orientation of one of the four "ideal" components of the texture, partly retains its orientation during annealing, while recovering from its cold-worked condition, and it is partially absorbed at the same time by invading strain-free grains of an orientation approximately corresponding to that of another "ideal" texture component. The reorientation here, as well as in the formation of the strain-free grains of "cube" orientation, may be described as a [Ill] rotation of about 40°, see Fig. 1 of ref. 6. The preferential growth of grains in such orientations is a result of the high mobility of grain boundaries corresponding to this relative orientation.' " It appears very likely that in copper the mechanism of the structural changes during annealing is similar to that observed in aluminum (except for the much greater frequency of formation of annealing twins in copper). In both metals the new grains of cube orientation have a great advantage over the new grains with orientations close to one of the four components of the rolling texture. This advantage stems from their symmetrical orientation with respect to all four retained rolling texture components of the matrix; they are oriented favorably for growth at the expense of all of these four orientations. As a result, the growth of the "cube grains" is favored over the growth of the others, as soon as the new grains have grown large enough to be in contact with portions of the matrix containing elements of more than one, and preferably of all four component textures. It is clear that this critical size is smaller and, therefore, attained earlier in the annealing process if the structural units, such as grains and kink bands, representing the four matrix orientations are smaller, i. e., if the average thickness of the rolled grains is smaller. Hence, for a given annealing time and temperature, a smaller penultimate grain size and a higher rolling reduction both tend to increase that fraction of the annealing period during which the above condition is satisfied. Consequently, the percentage volume of material assuming the cube orientation increases. The same is true also for increasing time and temperature of annealing when the penultimate grain size and the final rolling reduction are constant, since the average size attained by the new grains during annealing increases with the annealing time and temperature. For the same reason, at higher annealing temperatures a given volume percentage of cube texture can be obtained with larger rolled grain thickness (larger penultimate grain size, or smaller rolling reduction) than at lower annealing temperatures. The well-known conspicuous sharpness of the cube texture may be interpreted as a result of the fact that selective growth of only those grains is favored that have an orientation closely symmetrical with respect to all four components of the deformation texture and exhibit, therefore, a high boundary mobility in contact with each. The effect of alloying elements in suppressing the cube texture, as described by Dahl and Pawlek,' appears to be associated with a change in the rolling texture. For face-centered cubic metals, such as copper, which do exhibit the cube texture upon annealing, the rolling texture is always of the type described above, i. e., scattered around four "ideal orientations" of approximately (123) [112]. The addition of certain alloying elements, such as about 5 pct Zn or 0.05 pct P in copper, has the as yet unexplained effect of changing the rolling texture into the (110) 11121 type. This texture consists of two fairly sharply developed, twin related components. In such cases, as in 70-30 brass and in silver, the annealing texture again is related to the rolling texture by a [lll] rotation of about 30°, however, because of the different rolling texture to start from, it has no cube texture component. At higher temperatures, both in brassm and in silver," grain growth leads to a further change in texture: A [lll] rotation of the same amount, but in reversed direction, back to the original rolling texture.

Jan 1, 1952

By AIME AIME

THIS journal is fully indexed-every issue, as soon as published-in a set of card?, which thus brings ready to your hand any article on any subject in this journal or in any one of 1700 other journals in all languages. This Engineering Index Service, instituted in January, 1928, is to be found in the Public Libraries of Newark, Cleveland, Bridgeport, Baltimore and the John Crerar Library, Chicago. Engineers in any field who live or work in New Jersey will find this mammoth index to the latest engineering literature in the Public Library of Newark, N. J.

Jan 1, 1929

By AIME AIME

THE Employment Bureau of the F. A: E. S., conducted under- the direct supervision of the secretaries of the four Founder Societies, has wanted to extend its activities and usefulness but it is operating right up to the limit of its budget; consequently, anything which requires the expenditure of more funds cannot be -undertaken, for the present. A great deal of time has been spent in recording, classifying, and interviewing the men listed, but owing to the business depression, only a comparatively small number of men have been placed' luring the last few months. There are men in New York from all over the country. What the Bureau needs is more positions available. With this in view, a number of high-grade engineers, themselves in need of employment, volunteered their services for a publicity campaign in the City of New York. A definite mode of procedure was evolved, and the success of the movement has been considerable, so much so that it has been thought well to call it to the attention, of members elsewhere, so that similar committees ' and campaigns might be organized in other cities. Following is a chronological arrangement of the methods used.

Jan 1, 1921

By K. Harms, T. D. Jones

TO unload large tonnages of lead bullion cast in 100-lb. bars is a problem which has confronted the lead refineries for many years. The bars, on arrival, must be restacked for unloading by truck or hand-charged to charging machines. Selected crews of labor, capable of working at sustained speeds for five- or six-hour periods, are required for this exhausting task. Due to the fatigue involved crew personnel changes rapidly. During the war it became increasingly difficult to maintain crews, especially in critical labor areas, so that the development of casting and shipping ten-ton lead blocks from the smelters to the refineries became a must in our operating program. For many years lead refineries, when making interdepartment or intradepartment transfers, have cast lead blocks weighing five or ten tons which are easily handled by crane by merely attaching the crane hook to a steel anchor embedded in the frozen lead blocks. In previous years, the practice was to use regular smelter slag pots which, due to their rounded shape, produced what were called "buttons" by the refinery personnel. The button-type blocks were unsafe as they rolled and tipped easily and were unsatisfactory for stacking since no piling pattern could be followed. To remedy this situation, flat-bottom molds came into general use and later the bottoms were made concave to prevent slipping.

Jan 1, 1946

Charles H. Morgan, Worcester, Mass. (communication to the Secretary): In connection with Mr. Douglas's mention of the continuous charcoal kiln used at Kopparberg, some further data concerning this important invention may be of interest. Before giving these, however, I will say a word or two concerning other features of Swedish practice. In 1894 I spent two days at Soderfers with Mr. Tom Bergendal, the manager of the works, and was specially interested in the gas-producers and wood-drying kiln, illustrated in the paper contributed to our Transactions by Mr. Odelstjerna in that year.* The wood-drying kiln is, however, quite different in design and construction from Ljungberg's charcoal kiln. In the summer of 1894 it was discovered that by placing the gas-producer as close as possible to the open-hearth furnace, and working the producer hot, ordinary weather-dried wood, still containing a considerable percentage of moisture, could be used to advantage in the melting-furnace. Hence, the dryingkiln was needed for wet wood only. One of the iron-works of the Udeholm Go. was the first to adopt the Lundin saw-dust gas-producer and condenser, enabling them to make gas from saw-dust containing 45 per cent. of moisture, and convey the gas from the condenser through wooden pipes to the Siemens steel-melting furnace.† This sawdust gas gave very satisfactory results in iron-works; but the proprietors experienced much trouble by reason of the pollution of rivers by the tar-water from the condenser. On my visit, in 1894,I found that the Udeholm Co. had put in for their steel-melting furnaces the same wood-gas producers

Jan 1, 1899