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By J. W. MacDonald

Review of equipment, performance, and shaft design plans including head-frames for sinking and forms for lining are given. Also included are comparative data for shafts of 11 x20 ft and 193/4 x 30 ft with 756 and 666-ft depths, respectively. Advantage from curtain wall installation with "slip forms" which warrant consideration for normal lining are outlined. A preliminary review is also made of "chemical grouting" to stop water inflow leakage with Halliburton equipment and service. Two coal mine shafts recently completed in Franklin County, Ill., afford experience and accomplishment to warrant review. These shafts provide access to a tract with a potential recovery of one hundred million tons. This work provided the initial sinking by other than a contractor's organization since 1918. Four additional shafts required during the interim were sunk under contract arrangement with others. Supervisory contact and inspection afforded experience to govern planning and conduct of the recent work. TEST DRILLING Core test holes were drilled within each of the two shaft locations. Four additional cable tool holes were drilled on a radius of 700 ft from the shaft bottom to determine the coal seam elevations and govern the underground layout plans for each shaft. A graphic log of the hoist shaft test drillhole is shown by Fig. 1. A very soft sandstone was found at a depth of 16 ft. Cored recovery of the sandstone proved much weaker than the overlying compacted sand secured with coring facilities for soil recovery. Gradation from the very soft sandstone to fairly hard material occurred with the next 14 ft, extending to a depth of 30 ft. The weak character within the upper portion entailed need of adequate protection when sinking as described later. A part of the sandstone provided sufficient porosity and permeability to warrant further investigation. Adjacent cable tool drilling for temporary water supply disclosed inflow was limited to 2 gpm with a depth of 40 ft. It was increased to 12 gpm during the next 50 ft of drilling and occasioned subsequent precaution when sinking. General arrangement of facilities for sinking the hoist shaft is shown by Fig. 2. An available steel frame building was used to accommodate the hoist, compressor, shop, and shower or change room. The stiff leg derrick with 90-ft length of boom permitted removal of forms from the wood shop to the storage area or delivery within the shaft. Reinforcing steel and other needs were similarly handled, in addition to headframe erection. The headframe was placed over one end of the shaft to enable passage of forms for lining through the adjacent half next to the stiff leg derrick. The steel headframe as shown by Fig. 3 was fabricated locally. Side anchored cables from the

Jan 1, 1961

Phase equilibria have been determined experimentally for pseudo-ternary sections of the form ?MnO?-(CaO+MgO)-(SiO2+Al2O3) with a fixed Al2O3/SiO2 weight ratio of 0.17 and MgO/CaO weight ratios of 0.25 and 0.17 respectively for temperatures in the range 1473-1673 K. The primary phase fields present for the MgO/CaO weight ratio of 0.17 include manganosite (Mn,Mg,Ca)O; dicalcium silicate a-2(Ca,Mg,Mn)O?SiO2; merwinite 3CaO·(Mg,Mn)O·2SiO2; wollastonite [(Ca,Mg,Mn)O?SiO2]; diopside [(CaO,MgO,MnO,Al2O3)?SiO2]; tridymite (SiO2); tephroite [2(Mn,Mg)O?SiO2]; rhodonite [(Mn,Mg)O?SiO2] and melilite [2CaO?(MgO,MnO,Al2O3)?2(SiO2,Al2O3)]. For the section with MgO/CaO weight ratio of 0.25 the anorthite phase (CaO·Al2O3·2SiO2) is also present. The liquidus temperatures of ferro- and silico-manganese smelting slags have been determined. The liquidus temperatures at low MnO concentrations are shown to be principally dependent on the modified basicity weight ratio (CaO+MgO)/(SiO2+Al2O3).

Jan 1, 2004

By M. S. Smith

Paper written on project work carried out in partial fulfilment of B. Eng (Mining Engineering) degree The condition of tunnels at Premier has deteriorated to such an extent that an objective viewpoint on what leads to poor condition was required. Poor and untimely development results in schedule deviation, rehabilitation, and increased working costs. The project focused primarily on quality, and not on efficiency. The project gave management guidelines that need to be focused on in order to produce quality infrastructure. The conclusions were reached by doing literature searches, underground site investigations, benchmarking and computer simulations. From the results we can see that, in order to produce good quality tunnels, we have to have the basics working. This was found to be one of the major problems. Firstly, tunnel profiles are not to the design specification and no fixed blast designs existed. After revising tunnel profile statistics, over break increased to 10% (previously 7%) and development metres were 50 m short of target. Blasting was not monitored and designs not optimized for the conditions. Furthermore, geological factors are the main contributor to tunnel failure, although this process is initiated and aided by poor development (overbroken tunnels, incorrect profiles, direction and location) and blast damage. Continuing to develop tunnels in the current manner will cost R 1.5 million in additional costs for loading and supporting overbreak, whilst an additional 1 800 man-shifts will be required. Blast patterns were designed as a starting point for trials, while the implementation of a blast management system, development statistical database, tunnel hand-over procedures and a revision of the bonus system were proposed.

Jan 1, 2004

By Douglas D. Donald

The New Jersey Zinc Co. successfully holed through a 2 1/2-mile haulage tunnel connecting its new Ivanhoe shaft with the Van Mater Shaft at Austinville, Va. This 8x 10-ft cross-section tunnel was driven from both ends and met at a point approximately a mile from one shaft and a mile and a half from the other. In order to make certain that the faces met accurately on line and grade it was necessary to refine conventional mine surveying practices. The mines that are serviced by these shafts are located in the southwestern part of Virginia approximately midway between Roanoke and Bristol at the west foot of the Blue Ridge. While the presence of lead-zinc sulfide ores has been known in this area for more than 200 years, it was in 1902 that New Jersey Zinc Co. acquired the Austinville property, which has been in constant operation since that time. About ten years ago the company undertook an extensive diamond drilling program at Ivanhoe several miles to the southwest. This program was successful in locating sufficient ore to justify the purchase of a number of the properties in this area and plans were made to develop the new orebodies. The discovery of this ore as well as substantial addi- tional reserves at Austinville resulted in a decision by the company to increase its daily milling capacity to 3000 tons, of which about 1000 tons would be supplied from the Ivanhoe mine. After careful consideration of all possible methods of transporting this volume of ore from Ivanhoe to the mill, it became evident that diesel locomotive haulage through a 2½-mile tunnel between the two mines would best suit prevailing conditions. Since the lowest operating level, known as the 1100 level, of each mine would be at approximately the same elevation, it was obvious that such a tunnel should be driven on this level. Not only would this tunnel facilitate feeding Ivanhoe ore into the underground crusher and skip loading installation at Austinville, but it would also open up a known orebody 5000 ft southwest of the Van Mater shaft. In addition it would provide a valuable diamond drilling base from which a large and a potentially favorable but untested block of ground might be explored. In order to hasten the completion of the tunnel, it was felt that it should be driven from both ends. ENGINEERING CONTROL PROCEDURES It was vital that the tunnel junction be made accurately on line and grade; thus it was felt that engineering control of an unusually high degree of reliability was required to prevent an accumulation

Jan 1, 1959

By John C. Machisak

THIS CHAPTER of the Minerals Yearbook (Volume III) contains overall injury experience for coal mines, both anthracite and bituminous, coke plants, petroleum and natural gas, peat, and asphalt and related bitumens (native); metal mines, their metallurgical plants, including ore-dressing plants and primary nonferrous reduction plants and refineries; nonmetal mines; nonmetal mills; sand and gravel operations; slag (iron blast-furnace) plaints; and stone quarries and their related pants. Volume I of the yearbook contains a chapter showing injury experience and employment data treated separately by metal mines, nonmetal mines, quarries and their related plants, metallurgical plants (ore-dressing plants and primary nonferrous reduction plants and refineries are shown separately), nonmetal mills, sand and gravel plants, and iron blast-furnace slag plants. Volume II contains injury and employment experience in the fuel industries (coal, coke, petroleum and natural gas, asphalt and related bitumens (native), and peat. Injury and employment data are submitted voluntarily for the mineral operations through an industry canvass made by the Bureau of Mines, except for the coal industry, whose reports are mandatory. Information published in this chapter is the result of this canvass. Figures presented in this chapter for 1961 are preliminary except for anthracite, peat, coke, petroleum and natural gas, asphalt and related bitumens, and slag. These preliminary data, with the exception of bituminous coal, are a summation of reports received by May 31, 1962, from the operators showing their injury and employment experience. Estimations of injury and employment experience are not included in these preliminary data.

Jan 1, 1962

By V. Pawlowsky-Glahn, E. Pardo-Igúzquiza, J. J. Egozcue, R. A. Olea

Compositional data constitutes a special class of quantitative measurements involving parts of a whole. The sample space has an algebraic-geometric structure different from that of real-valued data. A subcomposition is a subset of all possible parts. When compositional data values include geographical locations, they are also regionalized variables. In the Earth sciences, geochemical analyses are a common form of regionalized compositional data. Ordinarily, there are measurements only at data locations. Geostatistics has proven to be the standard for spatial estimation of regionalized variables but, in general, the compositional character of the geochemical data has been ignored. This paper presents in detail an application of cokriging for the modelling of compositional data using a method that is consistent with the compositional character of the data. The uncertainty is evaluated by a Monte Carlo procedure. The method is illustrated for the contents of arsenic and iron in groundwaters in Bangladesh, which have the peculiarity of being measured in milligrams per litre, units for which the sum of all parts does not add to a constant. Practical results include maps of estimates of the geochemical elements in the original concentration units, as well as measures of uncertainty, such as the probability that the concentration may exceed a given threshold. Results indicate that probabilities of exceedance in previous studies of the same data are too low.

Jan 1, 2015

By Paul Hosford, Tim Haldane

"The paper describes the fast-track development of the project from feasibility study through to detailed design and construction, highlighting the process adaptations for handling sticky ore and water management in a tropical environment, and subsequent commissioning and operating experience of the mine. The paper discusses the geological setting, the environmental and socioeconomic considerations in developing and operating a mining operation in an underdeveloped country and the logistical challenges involved in bringing a large capital investment into Central America.INTRODUCTIONGlamis Gold Ltd. acquired the San Martin project in Honduras as a result of its merger with Mar-West Resources in late 1998. Despite the prevailing low gold price environment at the time of acquisition, Glamis anticipated rapid development of the San Martin project due to the quality of the deposit and other positive factors.Following favorable exploration and permitting activities in the first half of 1999, Glamis prepared a Feasibility study. In October 1999, Glamis decided to proceed with the development of the project and selected Hatch to provide the engineering, procurement and construction management services for the project. The San Martin project was in full-scale commercial operation approximately 14 months after the receipt of Board approval, and by July 2001 the San Martin project was operating at approximately 150% of its feasibility design capacity."

Jan 1, 2002

By J. A. Bartol

A survey was made of accident data and operating practices related to overloading and uneven load distribution on haulage trucks in noncoal open pit mines. Safety and economic aspects were analyzed. A bibliography of the literature was prepared. Review was made of available or projected systems for weighing or load indicating that might be applied to haulage trucks. A model specification was prepared for an overload detection system. Several promising systems for overload detection were identified. One was installed and tested briefly in a 170-ton rear dump truck in a copper surface mine.

Jan 1, 1982

By M J. Begbie, K B. Sporle

The Golden Cross low sulfidation epithermal deposit produced approximately 750,000 ounces of Au from two ore zones: the underground Empire Vein System and the open pit stockwork. The Empire Vein System has at least eight stages of vein formation: (1) sulfide veins, (2) black veins, (3) massive veins, (4) banded veins, (5) thin white veins, (6) crumbly veins, (7) massive/platy calcite veins, and (8) flattened rhomb calcite veins. Gold mineralisation occurred relatively early in the lifetime of the system, dominantly during 2 through 4. The late calcite veins (7 and 8) and early sulfide veins (1) are barren. The vein system formed in a dominantly extensional environment, though some compressional faulting may have also existed during vein formation. There is no evidence for significant strike-slip faulting during vein formation, highlighting the fact that explorationists should pursue on-site structural patterns rather than fashionable models.

Jan 1, 1998

By J. E. Fraley

Full-column anchored resin bolts were first used in underground mines in 1972. These bolts appear to improve support in many roof bolt applications, as indicated by the use of about 30 million resin bolts in 1980. In the mid-1970's, Bureau of Mines in-vestigators began looking for a substitute for resin because of its short supply and rapidly increasing price. The investigators selected accelerated gypsum plaster (CaSO. ·1/2H20), which becomes gypsum (CaSO.· 2H20) when mixed with water. Gypsum makes an effective alternate grout material because it hardens quickly, is readily available, and is inexpensive in comparison with other cementitious materials. The Bureau received a patent (U.s. patent 4,096,944) for a system combining water capsules (wax-encapsulated water) with plaster into cartridges resembling resin cartridges. This report presents the results of tests conducted by the Bureau on the plaster, the water capsules, and their combined use in cartridges. The tests were designed to determine how the gypsum-plaster water capsule bolting system performs and what happens in the hole while the bolt is being inserted through the cartridge. The test results indicate that water capsule plaster cartridges are a workable alternative to resin cartridges under normal or dry conditions.

Jan 1, 1987

By K. D. Hester

"1.0 INTRODUCTIONThe uranium industry started in Canada with the reopening of the gravity concentrator at Eldorado's Port Radium mine in the North West Territories in 1942. In 1948, uranium exploration and processing was opened to the private sector, and the Federal government offered special pricing incentives to the industry. Eldorado explored the Lake Athabaska deposits in northern Saskatchewan, and in 1953 opened the Beaverlodge mine. Other mines in the area soon followed. In 1953 also, the deposits in the Blind River - Elliot Lake area and in the Bancroft area of Ontario were discovered. Mines and mills were quickly developed in these areas.Canadian uranium production peaked at 15,900 short tons of U3O8 in 1959. In that year, uranium exports at $300,000,000 exceeded those of all other minerals in value. On October 1, 1959, Canadian uranium operations employed 13,900 people.Up to 1956, it was a seller's market. By 1958, world production of uranium was far ahead of requirements. Accordingly, in November of 1959, the USAEC announced it would not renew it's options for continued delivery of uranium from Canadian mines. This led to the ""stretchout"" of existing uranium contracts, and many mines were shut down (see Table 1) .In the early 1970's, a commercial uranium market seemed to be developing. In Ontario, at Elliot Lake and Bancroft, mines reopened. But competition was coming from the Athabaska basin of Saskatchewan. Currently it is these high grade, open pit deposits of northern Saskatchewan which are in the forefront of Canadian uranium developments. Canadian developments since 1942 reflect the political and market forces, as well as metallurgical progress and changes in environmental standards. This paper attempts to show these process changes in broad perspective."

Jan 1, 1987

By H. E. Bartlett

In the operation of some flotation plants treating platinum ores it was found that there were imbalances between the inputs and outputs. The inputs and outputs were sampled as slurries and it was determined that the imbalances arose from biases in the sampling, despite the samplers having been designed according to well-established rules. The launders feeding the sample cutters, and the sample cutters themselves were redesigned using finite element analysis to model the flows. A new set of design rules were formulated. The application of a Components of Variance Analysis, to determine the variance of primary and secondary sampling and of analysis in the lab for a concentrate is presented to illustrate the need for frequent sampling of process streams.

Jan 1, 2002

The changes in layout and design over 48 years, from the first crushing and shiploading plant installed at Whyalla in 1915, up to the present plant being built at Koolan Island, Yampi Sound may be summarized as follows.1. Smaller motors and lower power consumption have resulted froma. oil lubrication, where practicable, in place of grease,b. a high pressure "oil-lift" device to reduce starting torque in crushers with white metal bearings, andc. roller bearings in place of white metal.2. Short centre drives with pivot-based motors permit controlled loads on bearings and a reduction in size of crusher houses and motor rooms.3. Dust proofing and pressurising of crusher houses and motor rooms reduces working temperatures and improves cleanliness.4. Variable speed pan feeders give improved plant control, with the use of one man only, and remove the need for large surge bins between crushing stages.5. The use of diesel quarry trucks to dump ore through a "drive over"tip results in faster truck cycles.6. Elevated cylindrical silos with air operated chutes, replace long horizontal storage bins with multiple manually operated side discharge chutes for loading rail cars, and give faster and more accurate truck loading. The silos are charged by belt conveyors with automatic distributors.7. Increased shipping bin capacity (40,000 tons at Whyalla and 6O,OOO tons being built at Koolan Island) caters for the bigger vessels now in use and envisaged for the future. Automatic travelling distributors deliver ore to the bins in layers to ensure, maximum mixing and a more uniform grade for blast furnaces.INTRODUCTIONThe Broken Hill Proprietary Company Ltd has six major crushing plants in Australia at its principal sources of raw material. These are at Iron Knob bench G, Iron Knob scree ore plant, Iron Baron, and Cockatoo Island, all of which are ironstone operations. The other two plants are at Rapid Bay, a limestone quarry, and at Ardrossan, which is the Company's only source of dolomite.

Jan 1, 1963

By L. E. Mordberg, E. N. Nesterova

Geochemical studies of the bauxite deposits show that the primary composition is gibbsitic. Some trace elements were relatively mobile during bauxite formation with the content of Th, U and Zr increasing and that of Rb, Ni and Co decreasing. Y and Pb were relatively immobile during lateritisation; their contents and ratios were inherited from different types of underlying rocks. The supply of Ni and Co took place at the diagenetic or epigenetic stages; their ultimate source was the crystalline basement in the region. The studies support the view that the bauxite was likely to have originated as a result of in-situ lateritisation and not through sedimentary transport and redeposition

Jun 18, 1905

By L. Malmgren, A. Ansell

This paper deals with the influence from blast induced stress waves on the performance of shotcrete support in the Kiirunavaara mine. Time histories of stress waves from blasting were obtained from measurements during production blasts in the mine, followed by failure mapping. A finite element model, consisting of beam and spring elements, was used to calculate the shotcrete response to vibrations from blasting. The modelling approach was similar to that of a building during an earthquake, with accelerations measured in-situ used as loads. The analysis showed that the calculated adhesive stresses exceeded the strength in the interface between rock and shotcrete near to the blasted ring. It also indicated that tensile rock failure did occur. These types of failure will presumably result in ejection of joint defined rock blocks or rock slabs and will require a ductile support, such as fibre reinforced shotcrete. The predicted failure was confirmed by the mapping, which also showed that the zone of failed shotcrete was much larger in areas with plain shotcrete compared to areas with reinforced shotcrete. L’influence de l’onde de pression générée par les sautages à la mine de fer de Kiruna sur la performance du béton projeté est étudiée dans le présent article. Les ondes créées par les sautages ont été enregistrées et les domages résultants ont été cartographiés. La réponse du béton projeté aux vibrations de sautage est étudiée à l’aide d’un modèle d’éléments finis, constitué de poutres et de ressorts. L’approche est similaire à celle d’un immeuble subissant un tremblement de terre, avec les accélérations internes utilisées comme charges. L’analyse montre qu’à proximité des sautages, la contrainte adhésive dépasse la résistance de l’interface roche-béton. L’analyse confirme aussi l’existence de ruptures en tension. Ces types de rupture induisent probablement l’éjection des blocks définis par les structures du massif rocheux et requierent un support ductil, comme du béton projeté renforcé par exemple. Le type de rupture est confirmé par les observations de terrain, qui montrent aussi une extension plus importante des domages lorsque le béton projeté n’est pas renforcé. In diesem Konferenzbeitrag geht es um den Einfluss, der aus einer Sprengung hervorgerufenen Schwingungswellen auf die Ausführung der Spritzbetonverstärkung in dem Kiirunavaara Bergwerk. Die Aufzeichnungen der Schwingungswellen von Sprengungen gemäß einer Fehlerkartierung beziehen sich auf Messungen, die während der Sprengungen durchgeführt wurden. Ein Finite Elemente Modell, bestehend aus Stab- und Federelementen, wurde für die Berechnung der Spritzbetondeformation in Bezug auf die Schwingungen aus Sprengungen angewendet. Die Modellierungsmethode war ähnlich der eines Gebäudes während eines Erdbebens anhand der gemessenen Beschleunigungen vor Ort, die als Lasten benutzt wurden. Die Untersuchung hat gezeigt, dass die berechnete Haftspannung größer war als die Festigkeit in der Verbindung zwischen Gestein und Spritzbeton in der Nähe des Sprengungscentrums. Die Analyse hat gezeigt, dass die Zugfestigkeit des Gesteins versagte. Diese Arten von Versagen werden als Folge der Herausnahme von Gesteinsbrocken oder Gesteinsplatten erklärt und benötigen eine faserverstärkte Spritzbeton. Das vorausgesagte Versagen bestätigte sich beim Kartieren, welches auch zeigte, dass die Zone des versagenden Spritzbetons in Bereichen mit unverstärktem Spritzbeton viel größer war als in Bereichen mit armierten Spritzbeton.

Jan 1, 2003

By Richard Havens

Acidic and basic uranium leach plant residues located on the Navajo Indian Reservation, Tuba City, Ariz., were successfully stabilized against wind erosion using a relatively low-cost chemical method. An elastomeric polymer chemical was applied to the dike areas and a calcium magnesium lignosulfonate to the beach areas of three tailings ponds. The water-soluble chemicals were applied with an automated sprinkling system. The stabilization cost was $335 per acre for the 34.5-acre tract.

Jan 1, 1969

1. Scope of project In South Australia, the Western Mining Corporation operates the underground copper and uranium mine Olympic Dam. The deposits concerned are enormous (approx. 570 million tons) and have been mined since 1988. At the end of 1996, Western Mining decided to increase the capacity of Olympic Dam from 85,000 t.p.a. to 200,000 t.p.a. Planning At this point, SIEMAG-GHH Mining Division was asked to carry out a study including the planning of the entire shaft hoisting operations and calculation of the investment and operating costs involved. With the assistance of Australian partners, the study was completed within a single month. The principal points of the study were as follows: ? Planning of the hoisting machinery for 1,375 tons per hour from the surge bin below ground to the discharge bunker above ground. ? Optimisation of the hoisting operation with regard to maximum availability and safety. ? Optimisation of the time schedule to minimise the financing costs of the investment. ? Statement of the investment and operating costs. ? Comparison of the technical data, design and availability of the planned machinery with reference plants. Construction of the plant Following acceptance of the study, the decision was taken within a few weeks to implement it. SIEMAG Transplan GmbH received the order to execute the entire shaft hoisting machinery on the basis of the study submitted. The decisive factors in placing the order were: ? Convincing study, adherence to schedule ? Reference to comparable plants elsewhere ? Reference with regard to availability.

May 1, 2003

By M. H. Moys, J. A. Finch, A. R. Laplante

"A detailed review of ball milling circuit control practices in industry has been carried out. Based on fundamental process performance objectives, a standard system for base-level control of most ball mill installations is presented. This system is simple, inexpensive, and provides the major economic benefits available from contemporary automatic control technology.Value Analysis of Grinding OperationsValue analysis (or engineering, or management), originally developed by Lawrence D. Miles of the General Electric Company (1), focuses on the identification of value (product or process worth versus cost) and provides a systematic procedure for its improvement. A recent research and development project involving the application of the principles of value engineering to grinding operations has provided a basic technoeconomic model of grinding processes (2). The functional objectives of the ball milling process and the means to achieve value based grinding circuit control originate from this study.The Need for Grinding Circuit ControlThe economic effectiveness of a process is determined by the quality or value of the product and the related processing costs. If the inputs (both the raw feed and the applied processing work) are absolutely constant, the product characteristics and the process efficiency will be determined only by the engineering design. The need for and degree of process control is therefore determined by the nature of variations or upsets in the process inputs, and the extent to which these effect product value and operating costs."

Jan 1, 1988

An underground mine disaster is an accident of major proportions that takes a significant toll in human lives. A disaster usually disrupts the normal functioning of a mine and may result in entrapment of miners whose normal egress from the mine is cut off. A disaster often necessitates a rescue operation and a means of keeping the trapped miners alive while they await rescue. Explosion, fire, inundation of water or toxic gases, and collapse of a major portion of a mine are among the causes of mine disasters. Miners threatened by such an occurence must either bring the situation under control (e.g., by extinguishing a fire) or remove themselves from the danger. If control is unsuccessful, they must retreat to a place of safety. Four different activities may be involved: evacuation, escape, survival, and rescue. Evacuation is the orderly exit of people from the mine using previously identified escapeways and following a previously determined evacuation plan. Miners whose normal evacuation routes are blocked may be able to find alternative exit routes and escape by their own efforts. For those who cannot, the issue becomes survival while waiting to be rescued. The survival period may last for many days while rescue efforts are mounted from outside the mine.

Jan 1, 1981

This Bureau of Mines report describes research on methods of assuring the continued safety of mine electrical power distribution circuits through reliability and quality assessment of the protective devices, innovative inspection techniques for these devices and the circuits themselves, novel applications of trolley line circuit breakers, and computer modeling of distribution systems to determine the proper ratings for cables, power sources, and circuit protective elements. Battery charger safety is included.

Jan 1, 1979